BANCROFT 
LIBRARY 

o 

THE  LIBRARY 

OF 

THE  UNIVERSITY 
OF  CALIFORNIA 


E.  A.  HE.f.SAM, 
BERKELEY,  CAL 


The  Metallurgy 
of  the  Common  Metals 

Gold,  Silver,  Iron,  Copper,  Lead, 
and  Zinc 

by 


. 

Leonard  -S-.  Austin  >    j  «••+(* 

Professor  of  Metallurgy  and  Ore-dressing, 
Michigan  College  of  Mines. 


First  Edition 


1907 

Published  by  the 

Mining  and  Scientific  Press 

San  Francisco 


T"  N     6  6  tT 
.A? 


COPYRIGHT,  1907 

BY 
MINING  AND  SCIENTIFIC  PRESS 


6"  "7  7  7 


BANCROFT 
LIBRARY 


TABLE  OF  CONTENTS 


PAGE. 

PART     I  —  GENERAL 

Section  1.     Ores — Definition   and    Classification 13 

2.  Metallurgical  Treatment  of  Ores 16 

3.  Combustion     16 

4.  Fuels    20 

5.  Artificial  Fuels;    By-Product  Coke  Ovens;    Fuel  or  Pro- 

ducer   Gas 23 

6.  Refractory    Materials 32 

7.  Sampling ;    The  Sampling  of  Metals : 38 

8.  Preparatory  Breaking,  Grinding,  or  Crushing  of  Ores ...  48 

9.  Thermo-Chemistry  as  Applied  to  Metallurgy 55 

PART    II  — ROASTING 

Section  10.     The   Chemistry  of   Roasting 61 

11.  Roasting  of  Ores  in  Lump  Form 65 

12.  Roasting  Ores  in  Pulverized  Condition 72 

13.  Pot-Roasting  of  Ores 94 

14.  Cost   of   Roasting 94 

PART    III  — GOLD 

Section  15.     Gold  Ores 99 

16.  Gold-Milling   and   Amalgamation 99 

17.  Operation  of  a  Stamp  Battery 107 

18.  Leaching  Methods  for  the  Extraction  of  Gold  from- Ores  114 


TABLE    OF    CONTENTS 

PAGE 

19.  Chlorination  of  Gold  Ores 115 

20.  The  Vat  or  Plattner  Process  of  Chlorination 116 

21.  Barrel  Chlorination    .' 119 

22.  Cyaniding  of  Gold    (and   Silver)    Ores — General   Obser- 

vations      126 

23.  Development  of  the  Cyanide  Process 127 

24.  Chemistry  of  the   Cyanide   Process 129 

25.  Operation   of   Plant 130 

26.  Classification  of  Methods    134 

27.  First    Method    of    Cyaniding 135 

28.  Second  Method  of  Cyaniding.     South  Africa 144 

29.  Third    Method.      Mexico 148 

30.  Fourth  Method.     South  Dakota 156 

31.  Fifth   Method.     West   Australia 162 

32.  The    Bromo-Cyanogen    Process 164 

33.  Practice  upon  Cripple  Creek  Ores 166 

34.  Cyaniding    Sulpho-Telluride    Ores 168 

35.  Action  of  Copper  in   Cyaniding 168 

36.  Precipitation  with   Zinc-Dust 170 

37.  Extraction  of  Gold  by  Smelting 170 

PART    IV  — SILVER 
Section  38.     Silver  Ores   173 

39.  Silver  Milling  and  Amalgamation 174 

40.  Wet-Silver  Mill  or  Washoe  Process   (Tank  Settling)...   176 

41.  The  Boss  System  of  Silver  Milling 185 

42.  The  Combination  Process  of  Silver   Milling 186 

43.  Chloridizing    Roasting    190 

44.  Dry  Silver  Milling  (Reese  River  Process) 192 


TABLE    OF    CONTENTS 

PAGE 

45.  Hyposulphite  Lixiviation  of  Silver  Ores  (Patera  Process)  196 

46.  The  Russell  Process 201 

47.  The  Leaching  of  Argentiferous  Mattes 204 

48.  Cyaniding  Silver  Ores 209 

PART    V  — IRON 

Section  49.  Pig  Iron — Its  Manufacture  in  the  Blast- Furnace 213 

50.  Iron    Blast-Furnace    214 

51.  The  Stoves   219 

52.  Blast-Furnace    Plant    220 

53.  Operation  of  the  Furnace 220 

54.  Management  of  the  Iron  Blast-Furnace 223 

55.  Iron   Smelting    224 

56.  The  Chemical  Reactions  of  the  Iron  Blast-Furnace 225 

57.  Calculation  of  the  Iron-Blast-Furnace  Charge 228 

58.  Casting  the  Pig  Iron    230 

59.  Disposal    of   Cinder 231 

60.  Pig   Iron    231 

61.  Iron   Blast-Furnace   Slags 233 

PART    VI  — COPPER 
Section  62.     Ores    of    Copper 237 

63.  Smelting  the  Oxidized  Ores  of  Copper 240 

64.  Storage  of  Materials  of  the  Charge 246 

65.  Relative  Advantages  of  Blast-Furnace  and  Reverberatory 

Smelting    248 

66.  Calculation  of  Charge  in  Regular  Copper  Blast-Furnace 

Matting   249 

67.  Matte   and    Slag 252 


TABLE    OF    CONTENTS 

PAGE 

68.  Composition  of  Copper  Matte 253 

69.  The    Copper- Matting    Blast- Furnace 254 

70.  Blast-Furnace  Matte  Smelting  of  Copper 258 

71.  Regular    Matte    Smelting 258 

72.  Pyritic    Matte    Smelting 262 

73.  Pyrite  Smelting  at  Leadville,  Colorado 264 

74.  Matte  Concentration    264 

75.  Heat  Equations  of  the 'Copper- Matting  Furnace 265 

76.  The   Reverberatory   Matting  Furnace 266 

77.  Reverberatory   Matte    Smelting 268 

78.  Reverberatory     Smelting     of     Copper     (Welsh     Copper 

Process)     270 

79.  Direct  Treatment  of  Copper  Matte  for  Blister  Copper. .  272 

80.  Reverberatory   Matte-Smelting  at  Anaconda,   Montana..  273 

81.  Henderson   Process  for  the  Extraction  of  Copper  from 

Burnt   Pyrite    276 

82.  The  Hydrometallurgy  of  Copper 278 

83.  The    Copper    Converter 285 

84.  Converting  Copper  Matte  to  Blister  Copper 286 

85.  Re-lining  the  Converter  Vessels  or  Shells 289 

'86.     Accessories  of  a  Converter  Plant 292 

87.  Losses  in  Bessemerizing  or  Converting  Copper  Matte . . .  296 

88.  Cost    of    Converting 296 

89.  A  300-ton  Copper  Smelting  and  Refining  Plant 298 

PART    VII  — LEAD 
Section    90.     The    Ores    of   Lead 303 

91.  Bedding    Ores    304 

92.  Reverberatory   Lead   Smelting 306 


TABLE   OF    CONTENTS 

PAGE 

93.  The  Ore  Hearth 308 

94.  Silver-Lead   Smelting    310 

95.  The  Silver-Lead  Blast-Furnace 312 

96.  Blowing-In  the  Silver-Lead  Blast-Furnace 318 

97.  Reactions  of  the  Blast-Furnace 320 

98.  Lead  Slags    321 

99.  Calculation  of  a  Silver-Lead  Blast-Furnace  Charge 326 

100.  Handling    Base    Bullion 331 

101.  Flue-Dust  from  the  Silver-Lead  Blast-Furnace 331 

102.  Preparation  of  Flue-Dust  for  Re-Smelting 332 

103.  Treatment  of  Lead-Copper   Matte 334 

104.  Lead  Ores— Cost  of  Treatment 334 

PART    VIII  — ZINC 

Section  105.  Properties    of    Zinc 339 

106.  Ores  of  Zinc 339 

107.  Metallurgy   of   Zinc 340 

PART   IX  — REFINING 

Section  108.  Refining    of    Metals. 351 

109.  The  Parke's  Process  for  the  Refining  of  Base-Bullion..  352 

110.  Variations  in  Methods  of  Lead  Refining 362 

111.  Charges   for   Refining  Base-Bullion 363 

112.  Electrolytic  Refining  of   Copper 364 

113.  Costs  of  Electrolytic  Refining  of  Copper 366 

1 14.  Copper    Refining    Furnace 368 

115.  Copper   Refining    368 

116.  Parting    Gold-Silver    Bars .370 


TABLE    OF    CONTENTS 

PAGE 

PART   X  — COMMERCIAL 

Section  117.  Location  of  Reduction  Works 375 

1 18.  Handling    of    Materials 376 

119.  Organization  of  a  Metallurgical  Company 383 

120.  Investment  Required  on  Original  Plant 386 

121.  Profits     387 

122.  Organization     389 

123.  General  Remarks  on  Management  and  Labor 392 

124.  The  Purchasing  of  Ores  in  the  Rocky  Mountain  States.  401 

125.  The  Marketing  of  Ores  and  Metals 405 


PREFACE 


This  outline  of  the  metallurgy  of  the  common  metals,  namely, 
gold,  silver,  iron,  copper,  lead,  and  zinc,  is  devoted  to  the  de- 
scription of  the  processes  of  winning  the  metals  from  their  ores 
and  to  the  refining  of  these  metals,  except  iron,  the  metallurgy 
of  which  is  treated  only  to  the  point  where  pig  iron  is  obtained. 

Following  the  description  of  ores,  as  well  as  of  the  fuels  used 
in  treating  them,  and  the  materials  of  which  the  furnaces  are 
composed,  we  come  to  their  sampling,  for  the  determination  of 
their  exact  value  before  treatment. 

A  chapter  has  been  devoted  to  the  subject  of  thermo-chemistry 
as  applicable  to  igneous  methods  of  extraction. 

The  winning  or  reduction  of  the  various  metals  is  then  taken 
up  in  order  and  is  followed  by  a  description  of  the  methods  of 
refining  them. 

Attention  is  then  given  to  commercial  considerations,  since 
the  processes  must  be  conducted  in  a  profitable  way. 

The  author  is  indebted  to  Mr.  F.  L.  Bosqui,  who  has  not  only 
read  the  manuscript,  but  has  modified  the  portion  devoted  to 
the  cyaniding  of  gold  and  silver  ores,  as  his  special  knowledge 
has  justified. 

For  the  subject  matter  relating  to  the  smelting  of  silver-lead 
and  copper  ores,  the  author  has  drawn  on  his  own  experience, 
gained  during  upwards  of  a  quarter  of  a  century  of  practical 
work. 


PART  I.    GENERAL 


PART  I.  GENERAL. 

i.     ORES — DEFINITION  AND  CLASSIFICATION. 

An  ore  may  be  defined  as  a  mineral  aggregate  containing 
metals  in  quantity  sufficient  to  make  their  extraction  commercially 
profitable.  Minerals  or  rocks  containing  15  to  30%  Fe  would 
hardly  be  called  iron  ore,  nor  would  we  designate  as  an  ore,  silver 
ore  that  carried  but  2  or  3  oz.  silver  per  ton.  Nevertheless,  the 
rock  of  the  Treadwell  mines  on  Douglas  Island,  Alaska,  carrying 
$2.50  to  $3  in  gold  per  ton,  may  be  called  ore  since  it  can  be 
profitably  worked.  In  general,  ores  are  named  from  their  chief 
mineral  constituent,  as  a  lead,  a  copper,  or  a  silver  ore.  Such 
an  ore,  however,  may  contain  other  metals  than  those  which 
designate  it.  For  example,  a  lead  ore  may  carry  silver  and  gold, 
a  copper  ore  may  contain  silver,  gold  and  some  lead ;  a  silver 
ore  may  carry  lead  and  copper.  One  can  generally  tell  by 
appearance  whether  an  ore  carries  lead,  copper,  zinc  or  iron ;  but 
gold  and  silver  are  not  often  visible,  and  the  best  way  to 
determine  their  presence  is  by  assay. 

Ores  carrying  but  little  lead,  say  less  than  5%,  are  designated 
dry  ores.  Those  having  5  to  10%  Pb.  or  more  are  termed  leady 
ores.  The  former,  often  quite  silicious,  possess  value  chiefly  be- 
cause of  their  contained  gold  and  silver.  Copper  ores,  containing 
5  to  10%  copper,  and  free  from  lead,  also  contain  gold  and  silver 
in  appreciable  quantity. 

Mixed  ores  are  those  which  contain  both  lead  and  copper,  so 
that  it  is  often  puzzling  how  to  designate  them.  In  doubtful 
cases,  smelting  companies  have  purchased  such  ores  on  either 
the  basis  of  their  lead  or  copper  content,  under  the  plea  that 
in  extracting  one  of  these  metals  the  other  is  lost  or  wasted. 

Straight  silver  or  free-milling  silver  ores  are  free  from  lead 
and  copper  and  may  be  treated  by  amalgamation.  Likewise 
free-milling  gold  ores  are  those  containing  the  gold  in  metallic 
form,  and  are  amenable  to  treatment  by  amalgamation. 


14  THE    METALLURGY 

Lead-sih'er  or  Icad-silver-gold  ores. — These  ores  carry  lead 
in  such  quantity,  that  when  the  lead  is  reduced  by  smelting,  the 
precious  metals  alloy  with  it  and  are  recovered  with  the  lead. 

Copper-silver,  copper-silver-gold,  or  copper-gold  ores. — These 
ores,  when  smelted,  yield  up  their  copper  which  at  the  same  time 
takes  up  with  it  the  precious  metals. 

Base-metal  ores. — Lead  and  copper  ores  often  contain  zinc, 
antimony,  arsenic,  tellurium,  or  bismuth.  These,  in  the  process 
of  reduction,  often  alloy  themselves  with  the  principal  metal, 
to  its  commercial  detriment,  and  require  expensive  after-treat- 
ment for  their  removal. 

Plain  lead,  zinc  and  copper  ores. — These  contain  practically 
no  precious  metal  to  assist  in  defraying  the  cost  of  treatment. 
Consequently  they  must  be  of  higher  grade  to  be  profitably 
worked.  They  may,  however,  contain  small  amounts  of  silver 
and  gold,  but  not  enough  to  justify  treatment.  As  an  example, 
pig  lead  containing  4  oz.  silver  per  ton  would  not  pay  the  cost 
of  refining  for  the  purpose  of  extracting  silver ;  or,  blister  copper 
might  contain  as  much  as  16  oz.  silver  and  yet  not  pay  the  cost 
of  electrolytic  refining  for  its  recovery. 

Grading  ore. — Miners  often  find  it  profitable  to  sort  their  ore 
into  different  grades,  such  as  shipping  or  smelting  ore,  milling 
or  concentrating  ore,  according  to  the  after-treatment  which 
they  propose  to  give  it.  This  matter  is  one  of  the  most  important 
which  the  metallurgist  is  called  upon  to  consider  in  determining 
the  treatment  of  the  ore,  and  is  referred  to  later  in  speaking  of 
the  combination  method  for  the  treatment  of  a  mixed  silver  ore. 

Classification. — An  ore  consists  not  only  of  the  species  of  metal- 
lic compound  from  which  it  has  been  named,  but  also,  in  most 
instances,  of  gangue  or  waste  material.  On  the  other  hand  the 
ore  may  be  very  free  from  gangue,  as  in  the  case  of  galena,  zinc 
concentrate,  or  iron  ore;  or  an  ore  may  consist  mostly  of  waste 
material  with  comparatively  minute  quantities  of  the  metal  dis- 
seminated through  it,  as  a  gold  or  silver  ore. 

In  reference  to  their  metal  contents,  we  may  divide  ores  into : 

(i)  Straight  or  simple  ores,  namely,  gold,  silver,  copper,  lead, 
mercury,  zinc,  etc. 


OF    THE    COMMON    METALS.  15 

(2)  Complex  or  mixed  ores,  namely,  silver-gold,  silver-gold- 
lead,  lead-zinc-copper-silver. 

As  respects  impurities,  we  may  divide  ores  into  free-milling, 
refractory  or  rebellious,  docile,  arsenical,  antimonial.  A  free 
milling  ore  permits  the  extraction  of  most  of  its  gold  arid  silver 
by  the  simple  milling  operations  of  grinding  and  amalgamation. 
A  rebellious  ore  (a  term  formerly  more  used)  is  one  which  re- 
quires a  preliminary  treatment  by  roasting  before  it  can  be 
milled,  or  which  must  be  smelted.  Even  smelting  ores  may 
present  difficulties  which  would  entitle  them  to  be  called  re- 
bellious. A  docile  ore,  on  the  contrary,  is  one  which  may  be 
easily  treated.  Gold,  silver  or  copper  ores  carrying  arsenic  or 
antimony  may  be  called  refractory  or  rebellious. 

When  carrying  much  lead  or  zinc,  ores  are  designated  as  leady 
or  zincky.  When  free  from  lead  they  are  called  dry. 

As  respects  treatment,  we  may  have  free-milling,  leaching, 
chloridizing,  cyaniding,  or  smelting  ore. 

In  smelting  ores,  there  may  be  basic,  silicious,  dry,  coppery, 
leady,  or  fluorspar  ore. 

In  milling  and  leaching  there  may  be  talcose,  quartzose,  raw, 
roasting,  earthy,  argillaceous,  talcy,  light,  heavy,  or  base-metal 
ores. 

Among  iron  ores  are  bessemer  and  non-bessemer,  manganese, 
silicious,  phosphoritic,  basic,  and  open-hearth  ore. 

The  gangue,  or  waste  part  of  the  ore,  often  its  principal  con- 
stituent, may  be  earthy,  silicious,  argillaceous  or  clayey,  talcose, 
or  limy.  When,  as  is  often  the  case,  the  metal  is  in  the  heavy 
part  of  the  ore  and  the  lighter  part  is  gangue,  the  ore  may  be 
susceptible  of  concentration  or  dressing  to  remove  the  gangue. 
An  ore  capable  of  being  thus  treated  is  called  a  concentrating 
ore,  and  the  valuable  heavy  part  obtained  from  it  is  called 
concentrate. 

We  may  also  divide  ores  into  sulphide  and  oxidized.  As  a 
matter  of  fact,  these  varieties  merge  into  each  other,  and  it  often 
becomes  difficult  to  say  whether  the  ore  belongs  to  the  sulphide 
or  oxidized  variety. 


1 6  THE  METALLURGY 

2.     METALLURGICAL  TREATMENT  OF  ORES. 

In  winning  metals  from  their  ores  the  ideal  treatment  may  be 
divided  into  three  stages:  (i)  Preparation;  (2)  Extraction; 
(3)  Separation. 

While  these  three  stages  often  merge  into  one,  it  is  well  to 
consider  the  ideal  treatment  from  which  departures  can  be  made, 
or,  in  which  any  two  stages  may  be  united. 

(1)  Preparation. — This  has  reference  to  those  operations  by 
which  the  ore  is  fitted  to  undergo  the  later  ones.     It  is  by  sub- 
jecting the  ore  to  grinding,  breaking  or  comminuting,  or  to  roast- 
ing or  other  heat  preparation,  that  it  is  so  fitted. 

(2)  Extraction. — This  consists  in  having  present,  or  in  add- 
ing to  the  mineral  charge  or  mixture,  a  collector,  whose  duty 
shall  be  to  receive  or  take  up  from  the  ore  its  metal  content, 
bringing   it   often   into   a   much   smaller   bulk,    from   which   the 
metal  is  more  easily  parted  in  the  succeeding  operation.     The 
collector  or  carrier  may  be  one  of  the  metals  in  molten  form 
(mercury,  lead,  or  copper)  or  water  or  other  liquid. 

(3)  Separation. — This  consists  in  removing,  by  precipitation 
or  concentration,  the  metal  content  of  the  collector,  whereby  the 
metal  is  collected  by  itself  in  marketable  form.     This  does  not 
necessarily  mean  refining,  for  the  metal  may  contain  impurities 
and  still  be  in  marketable  form. 

3.     COMBUSTION. 

Fuels  such  as  wood,  coal,  charcoal,  and  coke  are  combustibles 
which  combine  with  the  oxygen  of  the  air  with  so  much  energy 
as  to  produce  light  and  heat.  Being  a  chemical  action,  it  can 
only  take  place  under  favoring  circumstances.  First,  there 
must  be  a  supply  of  air  coming  in  contact  with  the  fuel ;  second, 
there  must  be  plenty  of  surfa'ce  exposed  to  the  action  of  the  air 
if  we  desire  vigorous  combustion. 

A  jet  of  gas  will  not  ignite  when  a  poker  at  low  red  heat 
touches  it,  but  if  the  temperature  is  of  a  bright  red  the  gas  is 
inflamed.  Illuminating  gas  gives  a  large  flame,  while  charcoal 
and  coke  simply  glow.  Flame  then  is  burning  gas.  Soft  coal 
burns  with  a  flame  because  the  heat  of  burning  distils  the  gas  out 


OF    THE    COMMON    METALS. 


of  it,  while  the  residue  of  coke  gives  forth  a  short  flame  due  to  the 
burning  of  the  carbon  monoxide  coming  from  it. 

Hydrogen  burns  with  a  non-luminous  but  very  hot  flame. 
In  the  case  of  coal  gas,  if  the  air  supply  is  abundant  as  in  a 
Bunsen  burner,  the  flame  is  non-luminous,  while,  as  ordinarily 
burned,  it  is  luminous  because  of  the  presence  of  hydrocarbons. 
In  the  fire-box  of  a  reverberatory  furnace,  if  we  carry  a  thick  fire, 
say  1 8  to  24  in.  deep,  we  can  produce  a  long  flame  in  the  furnace. 
By  carrying  a  thin  fire,  sa.y  6  in.  deep  the  air  gets  through  the 
fuel  more  abundantly  and  the  flame  is  shorter ;  that  is,  the  hydro- 


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i«    0  •.•*.«,;£• 

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FIG  1.     WIND  FURNACE. 


FIG  2.     CUPOLA  FURNACE. 


carbons  are  speedily  consumed.  The  long  flame  is  desirable 
where  we  wish  to  produce  the  heat  farther  along  in  the  fur- 
nace. If  a  cool  surface  be  depressed  into  a  luminous  gas-flame, 
the  flame  is  cooled  and  soot  is  deposited  on  the  surface,  The 
combustion  of  carbon  is  therefore  prevented  and  smoke  is  pro- 
duced if  the  escaping  gases  from  a  soft-coal  fire  come  in  contact 
with  a  cool  surface,  or  again,  where  there  is  a  short  supply  of  air. 
As  an  illustration  of  the  phenomena  of  combustion  let  us  take 
the  case  of  a  fire  of  glowing  coke.  (Fig.  I.)  The  air  is  in 
excess  at  the  first  instant,  producing  carbon  dioxide  thus : 
(i)  C+20=C02 


1 8  THE    METALLURGY 

The  action  is  highly  exothermic,  or  heat-giving.  It  will  be 
noticed  that  the  nitrogen  is  present  as  a  dilutant  to  the  extent  of 
70%  of  the  total  gases  and  that  a  little  of  the  oxygen  manages  to 
escape  contact  with  the  coke,  especially  where  it  is  in  good-sized 
lumps.  At  2  to  4  in.  above  the  grate  in  a  clear  fire  we  may, 
therefore,  expect  to  find  the  greatest  heat.  But  as  we  go  upward 
this  condition  quickly  changes,  since,  in  presence  of  an  excess  gf 
glowing  carbon,  the  latter  is  dissolved  or  acted  on  by  the  carbon 
dioxide  as  follows : 

(2)  C02+C=2CO 

This  action  is  endothermic,  and  hence  we  may  expect  a  cool- 
ing zone  as  the  gases  rise  through  the  coke.  If  the  coke  is  in 
large  pieces  and  the  fire  thin,  a  little  air  may  draw  up  along  the 
walls,  and  escaping  combustion  and  issuing  above  the  fire  bed, 
unite  with  CO  gas,  burning  a  little  of  it  to  carbon  dioxide  thus : 

(3)   CO+O=CO2 

again  an  exothermic  reaction.     We  get  finally  a  mixture  of  gases 
much  as  follows: 

70%  N,  27%  CO,  2.5%  CO2,  0.5%  O,  called  producer  gas. 
Were  we  to  open  the  charging  door,  the  entering  air  would  burn 
the  CO  gas,  producing  a  blue  flickering  flame. 

To  obtain  the  greatest  amount  of  heat  from  the  fuel  the  thick- 
ness of  the  fire  should  be  no  greater  than  that  of  the  zone  (i,  i). 

Fig.  2  represents  the  conditions  in  a  blast-furnace  where,  as 
in  an  iron  foundry  cupola,  the  material  is  simply  mrelted.  The 
air,  forcing  itself  in  contact  with  the  glowing  coke,  as  per  the 
formula  (i),  rapidly  burns  it  with  the  production  of  an  intense 
heat  of  say  1500°  C,  this  action  continuing  to  the  upper  limits  of 
the  zone  (1,1)  where  the  oxygen  has  disappeared  and  where  the 
formation  of  CO  from  the  CO2  begins.  As  the  gases  enter  the 
colder  zone  (2,  2)  the  CO2  becomes  dissociated  in  the  presence  of 
carbon  to  CO,  but  little  of  it  remaining  at  the  upper  limits  of  that 
zone.  In  the  upper  zone  (3,  3)  the  cold  materials  of  the  charge, 
often  mixed  with  limestone,  tend  to  absorb  heat  from  the  rising 
gases.  The  limestone  is  dissociated  as  follows: 

CaO,  CO2=CaO+CO2=— 45500  calories. 

the  action  being  complete  at  800°   C.     In  a  cupola,  where  the 
operation  is  one  of  melting  only,  to  attain  the  greatest  economy  of 


OF   THE    COMMON    METALS.  IQ 

fuel,  the  coke  should  be  dense  and  in  large  pieces,  and  the  blowing 
should  be  done  with  plenty  of  air. 

From  the  equation  ( i )  we  shall  find,  that  to  burn  one  pound  of 
carbon  completely,  there  is  needed  2.66  Ib.  O,  or,  since  air  con- 
tains 23%  O  by  weight,  n.6  Ib.  air.  At  the  sea-level,  where  air 
weighs  12.4  Ib.  per  cu.  ft.,  this  will  make  143.8  cu.  ft.,  or,  in  round 
numbers,  150  cu.  ft.  per  pound  of  carbon.  Common  grades  of 
coke  contain  85%  of  carbon,  reducing  the  exact  figure  to  122 
cu.  ft.  to  one  pound  of  such  coke.  While  in  theory  12  Ib.  of  air  is 
sufficient  for  a  pound  of  coal,  it  has  been  found  that  the  air  is  not 
entirely  consumed  in  passing  through  the  fire,  and  that  an  excess 
should  be  used  to  ensure  complete  combustion.  With  a  forced 
under-grate  draft  16  Ib.  has  been  found  to  give  the  most  satis- 
factory results. 

The  quantity  of  heat  developed  in  unit  time  and,  therefore, 
its  intensity,  depends  upon  the  amount  of  coal  burned  in  that 
time  and  this  latter  upon  the  weight  of  air  supplied  to  it.  Now 
for  a  natural  draft  this  depends  upon  two  things,  the  grate-area, 
and  the  size  of  the  chimney  which  takes  away  the  products  of 
combustion  and  which  draws  the  air  through  the  fire.  For  a 
given  thickness  of  fire  and  draft-pressure  the  quantity  of  coal 
varies  directly  with  the  grate-area.  The  weight  of  air  passing 
away  by  the  chimney  is  the  product  of  its  velocity  into  its  weight 
per  cubic  foot  and  this  into  the  area  of  the  stack  thus :  W=Vaw ; 
where  W  is  the  weight  of  air  in  pounds  per  second,  V  the  velocity 
of  the  air  in  feet  per  second,  a  the  effective  sectional  area  of  the 
stack,  and  w  the  weight  of  a  cubic  foot  of  air.  The  velocity  V 
varies  with  the  square  root  of  the  height  of  the  chimney  and 
of  the  temperature  of  the  escaping  gases,  and  is  measured  by 
a  U-gauge  which  expresses  the  pressure  in  inches  of  water;  this 
pressure  varying  directly  with  the  height  of  the  stack,  and  conse- 
quently, as  the  square  root  of  the  velocity.  A  good  draft  is 
obtained  with  1.5  in.  of  water.  Since  the  velocity  depends  upon 
the  differences  of  temperature  between  the  column  of  air  filling 
the  chimney  and  that  of  the  outside  air,  we  have  the  formula : 


A- 


2O  THE    METALLURGY 

where  g  is  the  acceleration  due  to  gravity  (32.2  ft.  per  sec.),  h 
the  height  of  the  chimney,  T  the  internal  and  t  the  external  tem- 
perature in  degrees,  centigrade. 

Since  the  friction  of  the  chimney  and  of  the  escaping  gases 
comes  into  play,  the  effective  area  is  less  than  the  actual,  so  that 
we  may  give  the  effective  area  a=c — o.6^/e,  where  e  is  the  actual 
area. 

The  velocity  V  increases  as  the  square  root  of  the  increase 
of  temperature,  while  the  density  or  weight  per  cubic  foot  W  at 
the  same  time  decreases  at  the  rate  of  1-273  °f  its  volume  for 
each  degree  centigrade.  The  product  Vw  therefore  reaches  its 
maximum  when  the  chimney  gases  have  half  the  density  of  the 
outside  air  or  at  273°  C  (about  the  temperature  of  melting  lead), 
and  it  is  of  no  advantage  to  exceed  this  point.  (See  Fig.  3.)  At 
this  density,  25  cu.  ft.  of  air  equals  one  pound ;  this  multiplied  by 
24,  the  number  of  pounds  needed  in  practice  for  complete  com- 
bustion, means  600  cu.  ft.  per  pound  of  coal.  In  a  reverberatory 
melting  furnace,  where  a  neutral  flame  is  desired,  400  cu.  ft.  of 
air  is  sufficient.  On  the  other  hand,  in  a  roasting  furnace,  an 
oxidizing  atmosphere  may  be  needed,  in  which  case  air  is  sup- 
plied in  excess  of  that  required  for  burning  the  fuel,  entering 
the  furnace  by  holes  at  the  side  or  roof  or  through  the  bridge- wall. 

4.     FUELS. 

Fuels  may  be  divided  into  two  classes,  natural  and  artificial. 
Coal  and  wood  are  examples  of  a  natural  fuel ;  coke  and  charcoal, 
of  their  artificial  products. 

Xatural  fuels. — These  include  natural  gas,  the  mineral  oils,  and 
the  solid  fuels. 

Solid  natural  fuels. — These  may  be  divided  into  wood,  lignites, 
bituminous  coal  and  anthracite,  which  may  grade  into  one  an- 
other. This  is  particularly  true  of  the  latter  two.  We  may 
classify  them  by  composition,  making  the  fixed  carbon  and  vola- 
tile constituents  100%  and  excluding  ash,  sulphur  and  moisture, 
as  follows: 


OF   THE    COMMON    METALS. 


21 


Volatile 
Fixed  Carbon  Constituent 

Hard,   dry  anthracite 100  to  93%  o  to       7% 

Semi -anthracite    93  to  87%  7  to     13% 

Semi-bituminous    87  to  75%  13  to     25% 

Bituminous    75  to    *o%  25  to  100% 

Hard  or  anthracite  coal  may  be  considered  as  a  heavy,  com- 
pressed, natural  coke.     For  commercial  purposes,  it  is  customary 


O°  /O0°  2OO°  3OO*  400° 

FIG.   3.     TABLE   SHOWING  VARIATION   DUE  TO  TEMPERATURE. 

to  screen  it  into  different  sizes,  the  smaller  sizes  containing  the 
most  slate  and  dirt,  as  shown  by  this  table. 

Size  of  Coal.                                                            Fixed  Carbon  Ash 

Egg,  2.5  to  1.75  in 88.5^0  5-7% 

Stove,  1.75  to  1.25  in 83.7%  10.2% 

Chestnut,  1.25  to  0.75  in 80.7%  12.7% 

Pea,  0.75  to  0.50  in .79.0%  14-7% 

Buckwheat,  0.50  to  0.25  in 76.9%  16.6% 

A  cubic  foot  of  anthracite  coal  may  be  taken  at  60  to  65  Ib. 
and  its  calorific  power  at  6,500  to  7,000  calories.  The  prox- 
imate analysis  of  anthracite  coal  of  Pennsylvania  is  as  follows : 
3%  water,  3.4%  volatile  matter,  87.1%  fixed  carbon,  5.9%  ash, 
0.6%  sulphur.  Anthracite  coal,  having  little  or  no  flame,  is 
especially  suited  to  direct  contact  heating  as  in  the  blast-furnace 
or  in  the  crucible  melting  furnace. 

Bituminous  coals  are  chiefly  used  in  reverberatory  furnaces 
because  of  the  long  flame  which  they  produce.  As  an  example 


22  THE    METALLURGY 

we  may  take  West  Virginia  coal,  which  has  by  analysis  1.5% 
water,  37.8%  volatile  matter,  53.4%  fixed  carbon,  6%  ash,  1.3% 
sulphur,  and  having  a  calorific  power  of  7,333  calories.  We 
may  compare  this  with  an  example  of  semi-bituminous  Colorado 
coal,  containing  6.2%  water,  31.2%  volatile  matter,  52.5%  fixed 
carbon,  11.1%  ash;  or  with  a  Western  lignite,  as  that  of  Gallup, 
N.  M.,  which  contains  12.1%  water,  32.8%  volatile  matter,  47.6% 
fixed  carbon,  and  7.4%  ash.  Of  the  Rocky  Mountain  coals,  some 
will  coke,  but  the  larger  portion  will  not  fuse  in  the  coke-oven. 

Wood. — When  freshly  cut,  wood  will  contain  on  an  average 
40%  of  moisture,  and  where  it  can  be  burned  will  develop  2,300 
calories.  After  drying  out  for  several  months  it  still  retains 
20%  of  moisture,  and  its  calorific  value  has  increased  to  3,100 
calories.  Where  the  wood  is  perfectly  kiln-dried,  it  contains 
50%  carbon  and  has  a  calorific  power  ranging  from  3,667  cal. 
for  white  oak  to  5,546  cal.  for  long-leaf  pine.  In  outlying  dis- 
tricts of  the  West,  where  the  metallurgist  is  dependent  on  wood 
for  making  steam  or  for  metallurgical  operations,  the  accumu- 
lation of  a  supply  of  dry  wood  should  be  one  of  his  first  cares, 
since  in  no  way  will  his  forethought  be  better  rewarded.  He 
should  purchase  it  delivered  and  corded  ready  for  measurement, 
and  in  measuring  should  make  due  allowance  for  short  dimen- 
sions and  open  piling.  Cord-wood  should  cord  up  so  as  to  give 
70%  solid. 

Mineral  oil  or,  fuel  oil. — This  is  the  most  concentrated  of  fuels, 
and,  where  the  cost  justifies  its  adoption,  is  used  not  only  for 
steam-making  but  also  for  roasting  and  melting. 

Heating  power. — It  will  be  found,  in  burning  fuel  oil  from 
various  localities,  that  the  calorific  power  is  much  the  same  in 
all.  Beaumont,  Texas,  oil  has  a  calorific  power  of  10,820  calories, 
and  a  specific  gravity  of  0.88.  Oil  can  be  burned  so  as  to  give 
not  only  uniform  heat,  but  also  an  oxidizing  (as  in  roasting), 
or  reducing  flame,  as  may  be  desired.  The  air,  which  enters 
with  it,  should  be  pre-heated  as  well  as  the  oil,  and  it  will  be 
found  best  to  inject  it  under  high  steam  pressure.  Light  and 
heavy  oils  should  not  be  burned  in  mixture. 

In  Russia,  where  it  has  been  burned  in  open-hearth  steel  fur- 
naces of  10  to  15  tons  capacity,  the  oil  used  was  15  to  20%  of 


OF    THE    COMMON    METALS.  23 

the  charge.  As  regards  comparative  cost,  it  was  found  at  the 
Selby  Smelting  &  Lead  Works,  Vallejo  Junction,  Cal.,  that,  with 
oil  at  $1.71  per  bbl.  of  42  gal.,  and  coal  at  $6.00  per  ton,  the  saving 
in  using  oil  was  from  40  to  60%.  It  was  found  that  in  the  rever- 
beratory  matting-furnace  a  higher  or  lower  grade  of  matte  was 
produced  according  as  the  flame  was  more  or  less  oxidizing  or 
reducing. 

Natural  gas. — In  Ohio,  Indiana,  and  Kansas,  there  are  regions 
where  natural  gas  has  been  obtained  by  boring,  as  for  oil.  It 
is  the  most  efficient  of  gaseous  fuels,  having  a  calorific  power 
of  611  calories  per  cu.  ft.  or  27,861  calories  per  pound. 

5.     ARTIFICIAL  FUELS. 

These  may  be  divided  into  the  solid  prepared  fuels  and  fuel 
gas.  The  solid  fuels  are  charcoal  and  coke. 

Charcoal. — Wood,  packed  into  a  kiln  where  a  part  of  it  is 
permitted  to  burn,  heats  up  the  remainder  so  that  the  volatile 
portion  is  distilled,  leaving  charcoal.  The  latter  keeps  the  form 


FIG.  4.     SECTION  OF  CHARCOAL  KILN. 

of  the  wood  from  which  it  is  made,  and,  by  burning,  becomes 
much  lighter,  having  a  specific  gravity  of  0.2.  A  heaped  bushel 
weighs  14  to  1 6  Ib.  It  contains  95%  carbon  and  has  a  calorific 
power  of  7,610  calories. 

Charcoal  is  generally  made  in  a  kiln,  shown  in  section  in 
Fig.  4  and  in  perspective  in  Fig.  5.  It  is  set  at  the  foot  of  a 
steep  bank  so  that  it  can  be  conveniently  charged  from  above. 


24  THE    METALLURGY 

It  has  two  openings  A  and  B.  Wood  is  first  packed  in  at  the  lower 
opening  and  the  filling  finished  from  above.  There  are  three 
rows  of  3  by  4  in.  holes  set  2  ft.  apart  around  the  bottom.  The 
kiln  is  lighted  at  the  lower  door,  and,  when  fairly  started,  both 
openings  are  closed  with  sheet-iron  doors,  the  air  entering  by 
the  small  holes.  When  combustion  has  proceeded  far  enough, 
these  openings  are  closed,  and  the  kiln  is  allowed  to  cool.  The 
first  period  of  burning  lasts  eight  days,  the  cooling  four  days 
more.  Such  a  kiln  will  hold  25  cords  of  wood,  and  will  produce 
1125  bu.  of  charcoal  of  16  Ib.  per  bushel,  or  about  20%  of  the 
weight  of  the  wood  charged. 

The  ash  of  charcoal  amounts  to  from  I  to  3%,  that  of  the 
charcoal  made  from  wood  of  the  arid  regions  of  the  West  car- 
rying the  most.  The  ash  contains  but  little  sulphur  or  phos- 
phorus, and  hence  will  make,  in  the  iron  blast-furnace,  a  pure 
and  strong  pig  iron. 

The  Pierce  process. — This  is  intended  both  for  the  making  of 
charcoal  and  for  the  recovery  of  the  by-products  of  the 
destructive  distillation  of  wood.  The  wood  is  heated  in  brick  kilns 
by  burning  gas  arising  from  a  previous  operation,  air  being  at  the 
same  time  injected  by  means  of  steam- jets.  The  wood  is  dried 
out  in  about  18  hours  and  then  distillation  begins.  The  top  of  the 
kiln  is  then  closed,  and  the  products  drawn  away  to  condensers 
by  means  of  fans.  Part  of  the  uncondensed  gas  is,  however,  to- 
gether with  a  proper  proportion  of  air,  returned  to  the  kiln  to 
supply  additional  heat,  while  the  remainder  is  used  for  steam- 
making.  The  whole  operation  of  charging,  condensing,  cooling 
and  discharging  takes  8  days.  The  condensers  are  a  series  of 
copper  pipes  cooled  by  circulating  water.  The  wood  yields 

Charcoal    25.3  % 

Methyl  or  wood  alcohol 0.75 

Acetic  acid   i 

Tar    4 

Water    46 

Permanent  gases   23 

By-product  charcoal  is  quite  dense,  weighing  20  Ib.  per  bushel. 
By  the  sale  of  these  by-products,  and  because  of  the  superior 


OF    THE    COMMON    METALS.  25 

quality  of  iron  made  from  charcoal,  it  has  been  possible,  where 
the  supply  of  wood  was  abundant,  to  build  up  the  industry  of 
charcoal-iron  making  in  spite  of  the  competition  of  coke  blast- 
furnaces. 

Charcoal,  however,  is  more  friable  than  coke  and  makes  a 
good  deal  of  fines  or  'braize,'  as  it  is  called.  This  will  accu- 
mulate at  the  sides  of  the  iron  blast-furnace  tending  to  form 


FIG.  5.     PERSPECTIVE  VIEW  OF  CHARCOAL  KILN. 


'scaffolds.'  Even  in  the  lower  furnaces  for  lead  and  copper 
smelting,  a  charcoal,  made  from  a  light  wood,  is  apt  to  be  friable 
and  to  give  trouble  in  this  way,  and  only  the  firmer  woods 
should  be  used  in  charcoal-making. 

Coke. — This  is  made  in  kilns  in  a  similar  way  to  charcoal 
from  a  bituminous  coal  which  will  coke  or  fuse  together  at 
the  high  heat  of  the  kiln.  Coke  is  chiefly  made  in  bee-hive  ovens 
of  which  Fig.  6  is  a  type.  These  are  12  ft.  diam.  by  7  ft.  high  in 
the  clear.  They  are  charged  through  a  hole  in  the  roof,  each 
oven  holding  5  to  6  tons  of  the  finer  portion  of  the  coal  (less  than 
0.75  in.)  as  it  comes  from  the  mine.  In  burning  72-hour  coke 
the  charge  is,  in  the  morning,  dropped  into  the  hot  oven  from 
the  coal-larry  standing  above,  and  is  leveled  through  the  working 
opening  to  the  depth  of  26  in.  It  is  then  walled  up  with  brick, 


26  THE    METALLURGY 

an  opening  being  left  at  the  top  of  the  door  for  admission  of  air. 
Combustion  soon  begins,  and  the  usual  dark  smoke  escapes  by  the 
charge-opening.  After  four  hours  this  becomes  white,  the  coal 
ignites  or  'strikes,'  and  flames  issue  at  the  top.  For  twelve 
hours  the  oven  burns  with  a  dull  smoky  flame  above  the  sur- 
face of  the  charge.  On  the  second  day  the  flame  becomes  bright, 
and  the  air  supply  is  nearly  cut  off,  the  coking  being  completed 
in  55  hours.  The  flame  gradually  subsides,  the  whole  interior 
of  the  oven  being  red-hot.  The  oven  is  then  tightly  closed  and 
left  until  drawn  in  the  morning  of  the  fourth  day,  the  whole 
operation  taking  72  hours.  To  draw  the  coke,  the  brick-front  is 
taken  down  and  water  played  by  hose  into  the  oven.  After 
it  is  superficially  cooled,  the  coke  is  drawn  out  by  a  coke-drag 
having  a  long  handle.  As  it  is  withdrawn,  it  is  cooled  with  more 
water.  The  heating  has  taken  place  from  above  downward, 
leaving  the  coke  in  prismatic  masses  of  a  steel-gray  color. 

The  following  table  indicates  the  composition  of  Connellsville 
coal  and  the  coke  made  from  it. 

Coal  Coke 

Moisture  at   100°   C 1.2%  0.6% 

Volatile  matter .  31 .3  1.4 

Fixed  carbon    59.8  86. 

Ash    7.2  1 1. 1 

Sulphur    0.5  0.9 

Such  coal,  carefully  coked  in  a  modern  bee-hive  oven,  will 
yield  66%  of  marketable  coke.  Coke  can  also  be  made  in  48 
hours  but  the  72-hour  product  is  firmer.  The  efficiency  of  coke 
depends  upon  five  factors,  namely,  (i)  hardness  of  body,  (2) 
fully  developed  cell-structure,  (3)  purity,  (4)  uniform  quality 
and  (5)  coherence  in  handling.  Hardness  of  body  and  well- 
developed  cell-structure  are  apt  to  go  together,  the  latter  char- 
acteristic being  important  in  permitting  the  access  of  air  in 
combustion.  Purity  depends  upon  a  low  ash,  10%  being  good 
and  6  to  8%  exceptionally  pure.  The  sulphur  in  coke  intended 
for  iron  blast-furnace  work  should  be  under  i%,  a  good  coke 
containing  as  little  as  0.5  to  0.75^  sulphur.  For  lead  or  copper 
blast-furnaces  this  does  not  so  greatly  matter.  Uniform  quality 


OF   THE    COMMON    METALS.  2.J 

means  absence  of  black  ends,  which,  being  soft,  are  dissolved 
in  the  upper  part  of  the  furnace  by  carbon  dioxide  gas.  Coher- 
ence in  handling  is  evidently  important  where  the  coke  has  to 
be  transported  far  and  re-handled  at  the  smelting  works.  The 


LJ 
FIG.  6.    SECTIONS  OF  BY-PRODUCT  AND  BEE-HIVE  COKE  OVENS. 

coke-fines  are  injurious  to  the  action  of  the  furnace,  and  should 
be  kept  out  by  using  a  coke-fork. 

The  calorific  power  of  a  Pittsburg  coke,  containing  89%  fixed 
carbon,  10%  ash,  and  i%  sulphur,  may  be  given  at  7,272 
calories. 

BY-PRODUCT  COKE  OVENS. 

The  principal  ovens  of  this  kind  in  the  United  States  are  the 
Otto-Hoffman  and  the  Semet-Solway.  The  former  type  is  well 
shown  in  a  general  sectional  elevation  in  Fig.  7.  The  coking 
chamber  is  18  in.  wide,  33  ft.  long,  and  5  ft.  3  in.  high,  and  is 
closed  at  each  end  by  an  air-tight  cast-iron  door.  The  side-walls 
of  the  chamber  contain  flue-openings,  in  which  the  gas  evolved 
in  coking  is  burned  in  order  to  heat  the  coal.  There  may  be  20 
to  60  of  these  chambers  set  side  by  side  in  one  block.  The  oven 
being  hot  from  a  previous  charge,  the  doors  are  closed,  a  fresh 
charge  is  dropped  by  the  hoppers  and  leveled  off.  Distillation 
at  once  begins,  and  the  gases  are  led  away  to  condensing  cham- 
bers, where  they  are  freed  from  by-products — tar,  ammonia  and 
benzol — and  the  permanent  gases  returned  by  a  gas-main  to 


28  THE    METALLURGY, 

be  burned  around  the  ovens  in  the  flues  already  referred  to.  The 
products  of  combustion,  in  passing  away  to  the  stack,  go  through 
a  regenerating-chamber  containing  brick  checker-work,  while  air 
is  blown  through  the  other  chamber  at  the  other  side.  Thus  the 
gas  is  burned,  together  with  highly  heated  air,  producing  an  in- 
tense heat  in  the  walls  of  the  coking  chambers.  As  in  open- 
hearth  work,  this  action  is  then  reversed.  Upon  the  completion 
of  the  coking  in  24  to  36  hours,  the  end  doors  are  opened  and 
the  coke  pushed  out  by  means  of  the  ram  shown  at  the  left.  It 
is  then  sprinkled  with  water  and  loaded  into  cars  standing  on 
the  track  at  the  right.  The  yield  is  about  72%,  being  7%  more 
than  that  of  a  bee-hive  oven.  The  coke  is  harder,  denser,  but  more 
reliable  than  the  bee-hive  and  has  not  the  same  silvery  gloss 
which  we  see  in  Connellsville  coke. 

Costs. — The  actual  cost  of  making  coke  may  be  given  at  5oc. 
per  ton  in  bee-hive,  and  37c.  per  ton  in  by-product  ovens,  to 
which  must  be  added  the  cost  of  coal  to  make  it.  A  bee-hive 
plant  of  400  tons  daily  capacity,  operated  6  days  per  week  would 
cost  $60,000,  while  a  by-product  plant  would  figure  up  to 
$300,000. 

FUEL  OR  PRODUCER  GAS. 

The  use  of  gas  made  in  producers  is  extending,  and  has  many 
applications  in  generating  power  and  in  metallurgical  operations. 
It  is  therefore  proper  that  in  this  place  a  careful  study  should  be 
made  of  the  subject. 

We  show  in  Fig.  I  and  2,  in  the  section  on  combustion,  how 
producer-gas  is  formed  where  air  rises  through  a  thick  coke 
fire,  and  where  the  fuel  is  consequently  in  excess.  The  calorific 
power  of  one  pound  of  such  gas  would  be  simply  that  of  the 
CO  or 

Its  specific  heat  x  heat  of  combination      0.127x29000 

— : —  —  =  280  calories 

Atomic  weight  28 

By  blowing  steam  through  the  glowing  fire,  and  decomposing  it 
according  to  the  equation 

C+ H,O=:CO+2  H, 

we  produce  what  is  called  water  gas.  This  gas  has  a  much 
greater  heating  power  than  ordinary  producer  gas  on  account  of 


OF    THE    COMMON    METALS. 


29 


3O  THE    METALLURGY 

the  calorific  power  of  the  hydrogen,  one  pound  of  which  would 
give  29,100  calories  if  again  burned  to  water.  The  reaction 
by  which  water  gas  is  produced  is  highly  endothermic,  and  in 
consequence  of  its  rapidly  cooling  the  fire,  must  be  stopped  in 
a  few  minutes,  so  that  the  quantity  which  can  be  thus  made  is 
limited. 

A  great  variety  of  gas  producers  have  been  made ;  among 
the  more  recent  types  is : 

The  Taylor  revolving-bottom  producer. — This  consists  of  a 
steel  brick-lined  shell,  6  ft.  internal  diameter  (Fig.  8),  with 
an  iron  bosh  below  where  the  ashes  accumulate.  These  are  held 
up  by  a  grate  which  about  once  in  six  hours  is  revolved  in  order 
to  grind  up  and  remove  the  ashes.  A  central  pipe  is  to  be 
seen  rising  through  the  ashes,  and  by  it  both  the  air  and  steam 
may  be  introduced  to  the  fire.  At  the  left  may  be  seen  a  steam 
injector  by  which  both  steam  and  air  are  blown  into  the  fire. 
Peep  holes  at  the  side  show  its  condition  and  the  height  of  ashes. 
The  coal  is  charged  by  hopper  and  bell,  being  kept  closed  except 
at  the  instant  of  charging.  The  gases  are  led  off  at  a  large  exit 
pipe  near  the  top. 

Fig.  9  shows  a  complete  plant  of  the  Loomis-Pettibone  system, 
having  a  positive  blast  exhauster,  and  making  both  producer  and 
water  gas,  and  cooling  and  cleaning  it  for  use.  The  operation  is 
as  follows :  Assuming  that  there  are  hot  fires  in  both  of  the  gene- 
rators (or  producers),  the  air  is  drawn  through  generator  i,  and, 
in  so  doing,  burns  the  fuel  and  makes  producer  gas.  This  gene- 
rator may  have  just  received  fresh  coal,  and  the  coal  smoke,  tarry 
matter  and  producer  gas  from  it  are  together  drawn  through  the 
hot  fire  in  generator  2,  being  completely  burned  in  so  doing.  The 
gas  goes  off  through  the  valve  B  to  the  boiler  where  its  heat  is 
absorbed.  It  then  passes  over  to. a  scrubber  where  it  rises  through 
coke  set  on  perforated  trays  in  a  tower.  This  coke  is  continually 
sprinkled  with  water  sprays  so  that  the  gas  is  cooled  and 
cleaned.  Rising  to  the  top  of  the  tower,  it  filters  through  a  layer 
of  fine  shavings  or  'excelsior,'  which  removes  any  particles  of 
dust,  and  passing  on  goes  through  the  Root  positive-blast  ex- 
hauster shown  at  the  right,  and  finally,  by  the  pipe  Z,  to  the 
producer-gas  gasometer.  The  fire  in  generator  I,  having  be- 


OF    THE    COMMON    METALS. 


FIG.   8.     TAYLOR   REVOLVING   BOTTOM    PRODUCER. 


32  THE    METALLURGY 

come  clear  and  hot,  fresh  coal  is  put  into  generator  2  and  the 
valves  change  to  take  the  air  at  the  bottom  of  it.  The  course  of 
the  gas  is  then  from  2  to  I  through  valve  A  (valve  B  having  been 
shut),  to  the  boiler,  thence  to  the  scrubber  and  exhauster  to  the 
gasometer.  In  making  water  gas,  steam  is  introduced  under  the 
grate  of  the  generator  which  has  the  upward  current,  and  while 
the  fire  is  hot.  The  water  gas  made  in  one  generator  is  perfected 
by  passing  down  through  the  other,  it  having  been  found,  that 
unless  this  is  done,  a  portion  of  the  hydrogen  again  reverts  to 
steam.  The  making  of  water  gas,  having  been  kept  up  for  a 
few  minutes,  so  cools  the  fire  that  steam  must  be  shut  off,  and  air 
again  substituted.  While  water  gas  is  being  made,  it  may  pass 
on  by  the  pipe  Z  to  the  gasometer,  or  Z  may  be  closed  and  Y 
opened,  permitting  it  to  go  to  the  water-gas  gasometer.  "Water 
gas  having  a  higher  calorific  power,  is  often  reserved  for  certain 
heating  operations  for  which  producer  gas  would  be  unsuited. 
X  is  a  'purge  pipe'  which  is  opened  when  starting  so  that  the 
air  in  the  system  may  be  expelled  before  turning  the  gas  into 
the  gasometer.  Steam  may  also  be  admitted,  together  with  air, 
above  the  fire  where  it  mingles  with  the  producer  gas  and  en- 
riches it. 

6.     REFRACTORY  MATERIALS. 

In  addition  to  ordinary  building  materials,  such  as  red  brick, 
iron,  or  stone  used  upon  the  exterior  portion  of  furnaces,  re- 
fractory bricks  and  materials  are  needed  for  the  interior,  where 
a  high  temperature  or  the  corrosive  or  scouring  action  of  molten 
metallic  oxides  or  slags  must  be  resisted.  These  materials  may 
be  either  in  natural  form,  as  sand,  or  in  artificial  form,  as  brick. 

Sand  is  used  for  repairing  or  fettling  the  interior  borders  of  a 
reverberatory  furnace,  forming  a  bank  \vhich  covers  up  or  pro- 
tects the  portion  eroded  by  the  scouring  action  of  the  charge.  The 
repair  is  executed  after  a  charge  is  withdrawn,  and  when  the 
sides  of  the  emptied  furnace  are  free  and^clear.  The  sand  is 
thrown  in  by  means  of  shovels,  or  set  in  place  by  long-handled 
paddles.  Sometimes  a  little  clayey  material  is  added  to  the  sand 
so  that  it  may  be  made  into  coherent  balls.  These  are  then 
thrown  against  the  side,  adhering  to  it  or  pressed  against  the 


OF   THE    COMMON    METALS. 


33 


34  THE    METALLURGY 

spot  to  be  repaired  by  means  of  the  paddle  just  referred  to. 
The  bottoms  of  such  furnaces  are  also  made  of  this  sand.  Gan- 
ister  or  lining  is  a  mixture  of  crushed  quartz  or  other  silicious 
rock,  to  which  there  has  been  added  just  enough  clayey  material 
to  make  it  cohere.  It  is  largely  used  for  lining  converters,  and  in 
this  case  a  very  silicious  ore  carrying  some  gold  and  silver  may 
be  used  instead  of  barren  quartz  rock.  Since  this  material  is 
all  eaten  or  scoured  away,  and  since  its  metal  contents  are  ab- 
sorbed by  the  matte  constituting  the  charge,  it  is  in  reality  a 
kind  of  ore-smelting  performed  without  additional  expense  of 
treatment.  For  stopping  a  tap-hole  of  a  furnace,  or  for  closing 
joints  or  other  openings,  clay,  or  a  mixture  of  sand  and  clay, 
mixed  into  a  stiff  mud  or  adobe  may  be  employed.  Clay  alone 
shrinks  in  drying  and  the  admixture  of  sand  prevents  this.  In  place 
of  the  sand,  coarsely  ground  fire-brick  (called  also  chamotte, 
grog,  or  cement)  may  be  used.  This  moist,  firmly  coherent,  and 
plastic  mass  is  molded  into  conical  plugs,  and  the  plugs  stuck 
upon  the  end  of  a  dolly  or  stopper-rod  and  pressed  into  the  tap 
opening,  thus  stopping  the  flow  of  slag.  It  is  also  used  for  stop- 
ping cracks,  and,  in  larger  masses,  for  closing  openings  or  aper- 
tures of  the  furnace. 

Refractory  materials  may  be  divided  into  three  classes: 

Acid,  as  silica-brick,  sand,  and  ganister. 

Neutral,  as  graphite,  chrome  iron,  fire-clay,  bone  ash,  and  car- 
bon brick. 

Basic,  as  dolomite  and  magnesite. 

The  choice  of  one  or  the  other  of  these  depends  on \  the  follow- 
ing considerations : 

Acid  materials  may  be  used  to  resist  the  scouring  or  corrosive 
action  of  acid  slags,  or,  being  very  infusible,  for  the  roofs  or 
arches  of  furnaces  subject  to  highest  heats. 

Xentral  materials  resist  well  the  action  of  neutral  slags  or  of 
metals,  and,  in  the  case  of  open-hearth  furnaces,  are  interposed 
as  a  layer  between  the  silica-brick  roof  and  the  basic-lined 
hearth. 

Basic  materials  are  used  where  the  slags  are  quite  basic.  Such 
slags  would  quickly  scour  or  corrode  a  silicious  lining. 


OF   THE    COMMON    METALS.  35 

It  will  be  noted  that  all  the  above  materials  have  not  only  spe- 
cial resistant  power,  but  are  also  very  infusible.  This  is  partic- 
ularly the  case  with  carbon,  which  in  the  absence  of  air  or  nitre, 
is  quite  permanent.  This  is  illustrated  in  the  carbon  filament  of 
an  incandescent  electric  light. 

It  is  the  brick  made  from  this  material  that  we  will  now  discuss. 

Silica-brick. — If  quartz  rock  or  a  sandstone,  containing  98% 
SiO2,  is  mixed  with  a  little  lime  paste  made  from  caustic  lime  and 
moistened  with  water,  it  will  cohere  sufficiently  so  that  it  can  be 
molded  into  bricks.  These  are  dried  in  a  drying  oven  or  room 
by  steam  heat,  and  then  charged  into  kilns,  where  they  are  burned 
or  heated  to  a  white  heat.  Fig.  12  represents  a  down-draft  kiln. 
It  is  a  dome-shaped  oven  15  to  25  ft.  diameter,  coal-fired  by 
means  of  fireplaces  set  in  the  exterior  wall.  The  flame 
rises  to  the  crown  of  the  arch,  and,  passing  downward  through 
the  brick  piled  in  open  order,  leaves  through  flues  at  the  floor  of 
the  kiln,  connected  with  the  adjoining  stack.  Silica-bricks  are 
burned  at  the  highest  heat  attained  in  such  an  oven.  In  use  they 
resist  the  highest  temperatures,  and  expand  when  heated  up. 
Consequently,  expansion  joints  or  openings  are  provided,  which 
close  when  a  furnace  is  in  use.  Otherwise,  the  tie  rods  which  bind 
the  furnace  may  be  slackened,  being  tightened  again  when  the 
furnace  is  closed  down  and  cooled  off.  The  linear  expansion  to 
white  heat  is  2.5  per  cent. 

Graphite  or  plumbago. — This  is  essentially  carbon  containing 
a  little  iron  and  other  gangue  as  an  impurity.  It  is  used,  in  mix- 
ture, for  melting-crucibles  or  for  retorts,  containing  50%  graphite, 
45%  air-dried  clay  and  5%  sand.  The  graphite  prevents  shrink- 
ing and  cracking  and  adds  to  the  refractoriness  of  the  vessel. 

Chromite,  or  chrome-iron,  is  a  double  oxide  of  iron  and  chro- 
mium (Cr2O3  FeO).  For  making  brick  the  ore  should  have  not 
less  than  40%  Cr2O3  and  not  more  than  6%  SiO2.  It  is  made  into 
brick  by  crushing  it,  mixing  with  a  little  lime  as  in  the  case  of 
silica-brick,  and  burning  in  kilns.  Chromite  is  not  acted  on  by 
silicious  fluxes. 

Fire-brick  and  tile. — We  come  now  to  that  most  used  of  refrac- 
tory materials  made  from  fire-clay.  Clay,  in  its  purest  form  as 
kaolin,  contains  silica  and  alumina  in  such  proportions  as  to  make 


30  THE    METALLURGY 

it  difficultly  fusible.  The  presence  of  certain  bases,  even  in  small 
quantity,  may  much  increase  its  fusibility.  Clays,  nearly  free 
from  such  impurities,  and  therefore  quite  infusible,  together  with 
uncombined  silica,  form  fire-clay.  Thus,  wre  have  a  typical  air- 
dried  clay  of  the  following  analysis : 

Kaolin   (A1,O3  2  SiO,  2  H2O)    58.9% 

Uncombined  silica    ( SiCX  ) 36.1 

Impurities  (CaO,  FeO )   2.5 

Moisture  and  organic  matter   1.9 

It  is  the  combined  water  in  the  first  constituent  which  renders 
the  clay  plastic  when  pugged,  that  is,  when  after  grinding  it  is 
incorporated  with  water.  The  pure  clays  alone  are  unsuitable 
for  making  fire-brick,  since,  in  burning,  they  shrink  and  crack. 
It  is  necessary  therefore  to  mix  with  the  clay  some  non-shrinking 
substance,  such  as  sand,  ground  quartz,  or  coarsely  ground,  pre- 
viously burned,  clay  or  brick,  which  has  already  reached  its 
maximum  of  shrinkage.  This  latter  material  is  called  'chamotte' 
or  'grog.' 


FIG.  10.     BRICK  MOLD. 


The  manufacture  of  fire-brick. — The  fire-clay  to  be  used  is  first 
coarsely  ground  to  about  lo-mesh.     It  is  then  put  into  a  large 


OF    THE    COMMON    METALS. 


37 


square  pit  in  layers,  and  thoroughly  soaked  in  water  for  24  hours. 
From  the  pit  the  mixture  of  clay  and  grog  is  shoveled  into  a  ver- 
tical pug-mill  by  which  the  clay  is  thoroughly  incorporated.  There 
is  obtained  a  plastic  mass  ready  for  the  molding.  The  molding 
may  be  done  either  by  machine  or  by  hand.  In  the  latter  case, 
the  clay  is  made  in  balls  large  enough  to  fill  a  mold  (see  Fig.  10), 
and  thrown  into  it  with  all  the  force  possible.  It  fills  the  mold 
completely,  the  excess  being  cut  off  with  a  straight-edged  stick, 
and  the  bricks  dumped  out  on  a  pallet  or  board.  These  pallets 
are  placed  upon  racks  and  the  bricks  are  air-dried  until  they  are 
so  stiff  as  to  give  but  little  under  pressure  of  the  finger.  They 


FIG.    11.     BRICK   REPRESSING   MACHINE. 

are  then  put  through  a  re-pressing  machine  ( Fig.  1 1 ) ,  by  which 
they  take  their  exact  form.  When  pressed,  they  are  placed  again 
on  pallets,  set  on  brick-cars,  and  run  into  the  dryer  consisting  of 
chambers,  where,  under  the  action  of  steam  heat,  the  final  moisture 
is  removed.  The  bricks  are  now  piled  in  the  kiln  (Fig.  12),  where 
they  are  burned,  the  operation  taking  upward  of  two  or  three 
weeks. 

Fire-brick  varies  in  its  physical  characteristics  according  to 
composition,  the  more  aluminous  being  tougher  and  more  resist- 
ant to  molten  slag  or  matte,  while  the  silicious  is  more  infusible. 

Bone  ash  is  made  by  burning  bones  in  a  kiln  and  coarsely  grind- 
ing the  white  burned  residue.  Organic  matters  and  moisture  are 
both  removed,  and  an  impure  calcium  phosphate  is  left.  It  resists 
the  action  of  litharge,  and  is  accordingly  used  not  only  in  assaying 
but  for  forming  the  test  or  large  cupel  of  the  English  cupelling 
furnace. 


38  THE    METALLURGY 

Carbon  brick. — Gas-carbon,  such  as  is  used  for  arc-lights  in 
electric  lighting,  is  made  into  brick  with  a  limited  amount  of  gas- 
tar  and  burned  to  brick.  This  brick  has  been  found  to  be  quite 
resistant  in  a  reducing  atmosphere,  such  as  the  interior  walls  of 
a  blast-furnace. 

Both  dolomite  and  magnesite  are  basic  materials,  used  for 
lining  furnaces  where  basic  slags  are  formed,  such  as  would  attack 
or  scour  silicious  or  even  neutral  fire-brick.  Dolomite  is  a  mag- 
nesian  limestone,  and  the  more  magnesia  it  contains  the  better. 
Magnesite,  or  magnesium-oxide,  is  made  into  magnesia-brick 


FIG.   12.     SECTION   OF   BRICK  KILN. 

by  calcining  it  at  a  high  temperature.  It  is  usually  dark  brown 
from  the  presence  of  iron,  and  is  the  most  valuable  of  basic 
materials.  It  is  used  both  for  basic  converters  and  for  open- 
hearth  basic  furnaces  where  the  slag  contains  as  little  as  15% 
silica. 

7.    SAMPLING. 

The  first  step  necessary  in  the  treatment  of  an  ore  is  to  learn 
its  value  by  assaying.  As  a  preliminary  the  ore  must  be  sampled 
so  that  in  a  bulk  of  a  few  ounces  we  may  have  a  correct  rep- 


OF   THE    COMMON    METALS.  39 

resentative  of  the  entire  quantity  of  the  ore,  whether  it  be  a  few 
hundred  pounds  or  thousands  of  tons.  Moreover,  the  metals  re- 
sulting from  the  treatment,  especially  where  there  are  several 
together,  should  be  sampled  before  assaying  them  for  these 
various  metals.  Further,  the  ores  and  materials  used  in  treat- 
ment should  be  sampled  in  order  to  determine  their  exact  con- 
stitution and  so  learn  exactly  how  to  combine  and  to  treat  them. 
Ores  worth  thousands  of  dollars  are  often  bought  outright  from 
the  producer,  and  upon  their  correct  sampling  depends  the  amount 
paid ;  so  that  both  seller  and  buyer  are  equally  interested  in  hav- 
ing the  sampling  correctly  performed.  In  obtaining  the  value 
of  a  lot  of  ore  we  must  ascertain  its  weight  and  moisture,  and 
calculate  its  dry  weight ;  also  obtain  its  correct  sample  and  its 
correct  assay.  Thus,  suppose  we  have  a  lot  of  ore  weighing 
10,800  Ib.  with  7%  moisture  and  containing,  by  assay,  54%  lead 
worth  3c.  per  Ib.  Since  the  assay  is  made  upon  the  dry  weight, 
we  have,  after  deducting  the  moisture,  10,044  Ib-  of  ore  contain- 
ing 5,424  Ib.  of  lead,  worth,  at  3c.  per  pound  (neglecting  frac- 
tions), $162.72. 

Receiving  and  weighing. — Ores  may  come  into  a  reduction 
works  either  loose  or  in  sacks.  In  the  first  case,  whether  re- 
ceived in  wagons  or  in  cars,  the  weight  of  the  vehicle  and  ore 
together  is  taken  upon  the  scales  and  called  the  gross  weight. 
When  the  vehicle  (car  or  wagon)  is  emptied,  its  weight,  called 
the  tare,  is  taken ;  this,  subtracted  from  the  gross,  gives  us  the 
net  weight.  It  is  this  net  weight  which  we  have  above  called 
the  wet  weight.  When  the  ore  arrives  in  sacks  the  weight  may 
be  ascertained  in  the  same  way,  but  from  this  figure  must  be 
subtracted  the  weight  of  the  sacks.  Often  the  sacks  are  trans- 
ferred direct  to  the  scales,  weighed  there  and  afterward  emptied. 
The  sacks,  if  worth  it,  may  be  dried  and  returned  for  re-use. 
Sometimes  the  ore,  especially  when  frozen,  is  charged,  sack  and 
all,  into  the  blast-furnace,  the  sack  serving  to  hold  the  fine  con- 
tents together  until  smelted,  thus  preventing  loss  of  flue-dust. 
In  this  case  the  sample  is  made  up  from  small  portions  taken 
from  each  sack. 

The  moisture  sample. — In  theory  the  moisture  sample  should 
be  taken  at  the  instant  of  weighing,  since  obviously  the  ore  dries 


4O  THE    METALLURGY 

while  the  operation  is  delayed.  In  practice  the  sample  is  taken 
either  then  or  as  the  car  is  being  unloaded.  When  taken  from 
the  car,  a  hole  is  dug  in  the  ore  and  the  dried-out  surface  portion 
rejected.  The  remainder  is  put  into  a  covered  can  and  from 
it  50  oz.  are  taken  and  promptly  weighed  on  a  moisture  scale. 
After  cautious  drying  on  a  hot-plate  or,  better,  on  steam  coils, 
it  is  again  weighed,  the  loss  being  the  moisture. 

The  methods  of  sampling  may  be  divided  into  hand-sampling 
and  automatic  or  machine-sampling. 

Hand-sampling  includes  such  imperfect  methods  as  the  grab 
sample  and  the  trench  sample ;  while  among  the  complete  methods 
we  have  coning  and  quartering,  fractional  selection,  the  split 
shovel,  and  riffles. 

The  grab  sample. — This  consists  in  taking  evenly  from  dif- 
ferent places  on  the  pile,  lumps  and  handfuls  of  the  ore  to  be 
treated  further  as  follows :  the  lumps  are  broken  uniformly  to 
about  one-fourth  the  size  of  the  large  pieces,  and  the  whole 
(lumps  and  fine)  mixed  together  and  worked  down  by  one  of 
the  hand-sampling  methods  to  be  later  described.  The  hand- 
sample  is  used  as  a  quick  method  to  give  an  approximate  idea 
of  the  metals  in  the  pile  of  ore. 

The  trench  sample  is  taken  upon  a  large  pile  of  ore,  such  as 
a  dump,  cutting  transversely  through  it  by  means  of  a  trench. 
As  the  workman  proceeds  he  throws  the  bulk  or  principal  por- 
tion of  the  ore  to  one  side,  but  reserves  an  aliquot  portion  of  every 
tenth,  twentieth,  or  hundredth  shovelful  for  a  sample.  In  this 
way  the  ore  from  the  bottom  of  the  pile  is  represented,  which 
is  hardly  the  case  in  a  hand-sample.  As  a  variation  from  trench- 
ing, trial  pits  may  be  sunk  into  the  pile,  the  aliquot  portion  being 
reserved.  It  is  this  small,  reserved  part  of  the  whole  which  is 
subjected  to  regular  sampling  methods.  The  method  is  mani- 
festly imperfect,  since  most  of  the  pile  is  unrepresented. 

We  come  now  to  the  complete  methods  of  sampling,  one  of 
the  older  being : 

Coning  and  quartering. — In  connection  with  this  a  starting 
operation  is  to  shovel  over  the  entire  pile  taking  out  the 
aliquot  portion,  about  every  tenth  shovelful,  for  the  sam- 
ple. The  work,  to  save  unnecessary  labor,  is  generally  done 


OF   THE    COMMON    METALS.  4! 

as  the  car  or  wagon  is  being  unloaded.  The  sample  is  wheeled 
into  the  mill  and  put  through  a  rock-breaker  where  it  is  coarsely 
crushed  to  about  1.5  in.  size.  As  the  ore  is  wheeled  from  the 
crusher  it  is  put  into  a  circle  or  ring.  Standing  inside  this  ring 
the  men  slowly  circle  around  it,  shoveling  the  ore  to  the  apex 
of  a  cone  which  is  forming  in  the  centre.  The  ore  is  then 
worked  down  into  a  flat  disk.  This  is  marked  into  four  sectors 
(whence  the  name  of  quartering),  of  which  two  are  reserved 
and  two  rejected.  The  reserved  sectors  are  shoveled  back  into 
a  ring,  and  the  men  again  form  a  cone.  This  is  again  flattened 
down  and  quartered,  and  the  opposite  quarters  saved  out.  Thus 
the  process  goes  on,  coning  and  quartering  the  ore,  until  a  single 
maximum-sized  rich  piece  would  appreciably  raise  the  grade  of 
the  sample  if  retained  or  lower  it  if  rejected.  At  this  point  the 
ore  is  crushed  to  0.5  in.  size  by  means  of  rolls.  Coning  and  quar- 
tering is  resumed  until  crushing  finer,  say  to  wheat  size,  is  neces- 
sary. The  ore  is  worked  down  to  two  pounds  in  weight  and  again 
finely  ground  to  the  desired  finishing  mesh  of  80  to  the  inch.  It 
is  thoroughly  mixed  by  'rolling'  on  a  piece  of  rubber  cloth.  This 
final  product  is  placed  in  four  4-oz.  bottles  or  in  paper  sample- 
sacks  and  marked  with  the  name  and  particulars  of  the  lot  of  ore. 

Fractional  selection. — This  differs  from  the  quartering  method 
in  that  every  second,  or  every  fourth  shovelful  of  the  lot  is  re- 
tained for  sample,  being  made  into  a  cone  for  the  purpose  of 
mixing  it.  From  this  cone  every  second  or  fourth  shovelful  is 
again  removed,  and  so  on  until  it  is  necessary  to  re-crush  the 
ore.  After  crushing  the  reduction  takes  place  in  the  same  way 
until  the  bulk  has  been  made  as  small  as  desired. 

The  split  shovel. — By  this  method  a  shovel,  resembling  a  fork 
and  made  up  of  several  troughs,  is  used.  Half  of  the  ore  drops 
through  the  spaces  between  the  trough  and  half  is  retained  as 
a  sample.  The  sample  is  passed  over  the  split  shovel,  again 
reducing  it  by  one  half,  and  so  on,  until,  as  in  the  other  methods, 
the  entire  quantity  has  been  cut  down  to  the  desired  amount. 

Riffles. — This  apparatus  is  used  only  for  finishing  a  sample 
after  it  has  been  ground  to  0.5  in.  size  or  less.  A  riffle  is  a  series 
of  troughs  1.5  to  0.5  in.  wide  set  in  a  frame  and  with  equal 
spaces  between.  As  in  the  split  shovel,  one  half  the  ore  is  re- 


THE    METALLURGY 


tained  in  the  troughs  of  the  riffle  and  one  half  rejected.     It  has 

the  advantage  that  all  parts  of  each  shovelful  thrown  upon  it 

get  representation. 

Machine  sampling. — It  will  be  seen  that  the  different  methods 

of  hand-sampling  involve  much  labor,  and  it  has  been  sought  to 
overcome  this  by  the  use  of  machinery. 
Machines  are  of  two  different  kinds:  (i) 
Those  which  take  part  of  the  stream  of  ore 
all  the  time,  and  (2)  those  which  take  the 
whole  stream  at  frequent  intervals.  Since 
the  stream  of  ore  is  not  homogeneous  the 
first  method  is  defective  and  the  second  is 
accordingly  preferred.  In  the  second  method 
the  scoop  which  passes  through  the  stream 
should  enter  always  at  the  same  side  since 
otherwise  more  is  taken  of  the  entering  side 
which  again  may  be  richer  or  poorer  than 
the  leaving  side  of  the  stream.  Fig.  13  is 
an  example  of  the  first  type,  the  ore,  as  it 
falls,  having  portions  successively  cut  out 
until  at  the  bottom  a  small  sample  remains. 
The  remainder,  falling  outside  the  tube,  is 
rejected  and  is  called  the  'principal.' 

The  Yezin  sampler  (Fig.  14)  is  an  ex- 
ample of  the  second  kind.  It  consists  of  a 
sheet-iron  double  cone  carrying  scoops,  and 
revolving  30  times  per  minute.  A  feed- 
chute  delivers  a  stream  of  ore  into  the  main 
hopper  while  the  scoops,  as  they  revolve, 
take  out  a  portion  of  it  to  be  delivered  at 
the  apex  of  the  lower  cone.  The  scoops 
together  occupy  one-fifth  the  circumference 
of  the  circle,  and  accordingly,  take  out  one- 
fifth  of  the  ore.  One  of  them  may  be 
omitted,  so  that  one-tenth  may  be  retained. 
By  repeating  the  process  with  another  sam- 
pler, or  by  sending  it  back  to  the  first  one, 
FIG.  13.  the  sample  is  reduced  in  bulk.  For  finish- 

PIPE  ORE-SAMPLER,      ing,  a  riffle  is  used. 


OF    THE    COMMON    METALS. 


43 


FIG.  14.    VEZIN  SAMPLING  MACHINE. 


44 


THE    METALLURGY 


Fig.  15  and  16  represent  in  plan  and  elevation  a  small  sampling 
works  of  a  capacity  of  10  tons  hourly.  One-fifth  to  one-tenth 
of  the  ore  in  the  car  is  delivered  to  the  mill,  the  principal  being 
retained  in  the  car  for  delivery  to  the  main  storage  bins.  From 
the  Blake  crusher  the  ore  passes  by  elevator  to  a  Vezin  sampler, 
where  one-fifth  (or  2  tons)  is  cut  out  and  passes  on  to  the  Dodge 
crusher,  for  finer  crushing.  It  is  then  elevated  to  a  second  Vezin 
sampler  which  again  cuts  out  one-fifth,  leaving  800  Ib.  It  then 
passes  to  the  large  rolls.  It  is  now  elevated  to  a  third  Yezin 
sampler,  which  cuts  out  one-tenth,  leaving  80  Ib.  of  sample,  this 
passing  on  to  the  small  rolls.  This  quantity,  i-25oth  of  the  ore, 
may  be  reduced  by  quartering,  and  the  small  amount  left,  ground 
in  a  sample  grinder.  The  discarded  portions  of  the  ore  are  con- 
ducted by  means  of  an  elevator  and  swinging  spout  into  any  de- 
sired bin  where  it  is  retained  until  the  ore  has  been  settled  for. 

Requirements  for  accurate  sampling: 

(i).  Frequent  taking,  which  will  ensure  representation  of 
every  part  of  the  ore. 

(2).  Mixing  thoroughly,  which  makes  the  ore  uniform 
throughout. 

In  progressive  crushing  we  make  the  ore  finer  as  the  sample 
diminishes  in  weight,  so  that  the  ratio  which  the  size  a  single 
rich  piece  bears  to  the  size  of  the  sample  shall  not  be  exceeded. 
The  following  table  expresses  these  relations : 

Table  showing  size  to  which  ore  is  to  be  crushed. 

Value  in  Silver  Oz.  per  Ton 


Weight  of  Ore 

Highest  300 
Average  50 

Highest  3,000 

Average  75 

Highest  10,000 
Average  500 

100  tons  to  10  tons  

Cocoanut 

Fist 

Fist 

10  tons  to  1  ton  

Orange 

Egg 

Walnut 

1  ton  to  200  Ib  

Walnut 

Chestnut 

Chestnut 

200  Ib  to  5  Ib 

Pea 

Wheat 

Wheat 

5  Ib.  to  bottle  sample.  .  .  . 
Bottle    samole.. 

20-Mesh 
80-Mesh 

25-Mesh 
100-Mesh 

50-Mesh 
120-  Mesh 

Cost  of  sampling. — The  cost  of  handling  the  ore,  delivering  it 
to  bins,  and  wheeling  the  retained  fifth  or  tenth  into  the  sampling 
mill  may  be  taken  at  loc.  per  ton.  The  cost  of  hand-sampling 


OF   THE    COMMON    METALS. 


45 


THE    METALLURGY 


this  retained  portion  will  be  750.  per  ton.  At  the  Metallic  Ex- 
traction Co/s  Works,  Cyanide,  Colo.,  ore  was  thrown  out  of 
the  car  into  a  feed  chute,  and  was  coarse-crushed,  automatically 
sampled,  and  delivered  to  storage  bins  for  nc.  per  ton.  At  cus- 
tom sampling  mills  a  charge  of  $i  to  $2  per  ton  has  been  made 
for  sampling,  storing  and  selling  the  ore,  the  sampling  mill  acting 
as  selling  agent.  This  price  is  made  even  lower  for  concentrates 
requiring  no  crushing. 

THE  SAMPLING  OF  METALS. 

Bars  of  gold  or  silver  are  sampled  for  assay,  either  by  taking 
granulations,  or  by  a  chip-sample.  In  the  first  case,  while  the 
metal  is  in  a  molten  condition,  a  small  ladleful,  of  an  ounce  or 
less,  is  taken  and  poured  into  a  bucketful  of  water.  This  gran- 


^602 

260.6 

2641 

269.0 

~l 
2609 

\ 

264.6 

250.0 

249.0 

249.0 

/ 

264.6 

265.5 
\ 

2570 

242.0 

258.0 

~1 

263.5 
1 

\ 

26/.0 

259.6 

26/.5 

/ 

Average  258.2oz.  stiver  per  Ton 

FIG.   17.     SECTION  OF  BAR  OF  BASE-BULLION . 

ulates  the  metal,  so  that  it  can  be  easily  weighed  in  assaying. 
The  chip-sample  is  taken  from  diagonally  opposite  edges  of  the 
bar  or  ingot  by  cutting  with  a  cold-chisel.  The  average  of  the 
two  chips,  thus  obtained,  is  taken  for  the  true  result. 

Base  bullion. — This  is  the  lead  which  comes  from  the  silver- 
lead  blast-furnace,  containing  commonly  100  to  500  oz.  silver  per 
ton.  When  poured  into  molds  and  allowed  to  cool,  it  will  segre- 
gate ;  that  is,  the  exterior  or  first-cooled  portions  of  the  bar  will  be 
higher  by  several  ounces  than  the  central  portion,  as  shown  in 


OF    THE    COMMON    METALS.  47 

cross-section  Fig.  17.  Base  bullion  was  formerly  sampled  by 
taking  two  chips  or  punchings,  one  from  the  top  and  one  from 
the  bottom  of  each  bar.  The  punch,  Fig.  18  (somewhat  resem- 
bling a  belt  punch),  1.125  in.  diam.,  was  driven  into  the  soft  metal 
by  a  sledge,  taking  a  cylindrical  piece.  In  a  carload  of  400  bars 
there  would  therefore  be  800  chips.  These  chips,  melted  together, 


FIG.  18.     BASE-BULLION  SAMPLING  PUNCH. 

would  be  the  sample  of  the  4OO-bar  lot.  The  preferred  way, 
though,  is  to  re-melt  the  metal  in  large  kettles,  skimming  and 
stirring  it,  and  casting  it  again  into  bars  or  ingots.  While  mold- 
ing, portions  are  taken  from  the  molten  bath  and  poured  into 
bullet  molds,  each  bullet  being  approximately  one-half  assay  ton, 
which  is  then  trimmed  to  the  exact  weight. 


73.9  2O/.  /  ///  6  J00. 7\  36. 6 
\  0.34\  0.34\O.24\  O.22\  O.22 


69./ 
\  O.2.6\0.2O 

\672 


\  7/.3\  70.3?\  63.8\ 


0.2.2\  0.2Z\  O.  Z2j  O.2.2 


FIG.  19.     SECTION  OF  BAR  OF  INGOT  COPPER. 

The  segregation  of  copper,  in  bars,  is  much  more  marked  as 
shown  in  Fig.  19,  which  gives  the  values  in  ounces  per  ton  of 
both  the  contained  silver  and  gold  in  a  pig  of  blister  copper  5 
in.  deep.  The  usual  method  of  sampling  such  bars  is  to  drill 
them,  retaining  the  shavings  or  chips  for  a  sample.  Manifestly, 
such  a  sample  would  be  uncertain,  depending  upon  where  it  was 
taken.  To  minimize  this  difficulty,  it  is  customary,  in  a  given 


48  THE    METALLURGY 

lot,  to  drill  each  succeeding  bar  in  a  different  spot  so  as  to  obtain 
an  average.  The  quality  of  copper  for  conductivity  is  often  a 
factor,  for  which  a  dipped  sample  of  the  charge  of  copper  in 
the  furnace,  just  before  it  is  molded,  is  made  into  a  wire  and 
tested. 

In  sampling  base-bullion  where  the  variations  are  smaller,  chip 
samples  are  fairly  accurate,  but  not  in  anode  or  blister  copper. 
To  ensure  such  accuracy  it  is  now  becoming  the  custom  to  melt 
the  whole  lot  of  metal  to  be  sampled,  to  stir  it  well  together,  to 
take  out  a  ladleful  and  to  pour  it  into  water  in  the  case  of  copper, 
or  into  a  small  mold  in  the  case  of  lead,  the  idea  in  the  latter 
case  being  to  cool  the  metal  so  quickly  that  segregation  can- 
not occur. 

Pig  iron  is  sampled  and  graded  both  by  its  appearance  when 
fractured  and  also  by  analysis.  When  the  latter  method  is  used, 
the  determination  is  made  from  drillings  taken  from  a  small 
molded  sample.  The  silicon  present  is  the  chief  determining 
factor  of  its  grade. 

The  sampling  of  metals  is  described  at  greater  length  in  dis- 
cussing their  reduction  from  ores. 

8.     PREPARATORY  BREAKING,  GRINDING,  OR  CRUSHING  OF  ORES. 

The  amount  and  method  of  breaking  or  crushing  varies  ac- 
cording to  the  further  treatment  of  the  ore. 

Ore  frequently  comes  from  the  mine  or  the  concentrator 
already  hand-sorted,  and  from  the  concentrating  mill  more  or 
less  finely  ground,  in  which  case  it  is  already  fine  enough  for 
metallurgical  treatment  by  the  several  methods  to  be  described. 

Crushing  for  sampling. — This  operation  varies  with  the  facil- 
ities for  crushing,  and  with  the  nature  of  the  ore  and  the  coarse- 
ness at  which  it  is  desired  to  keep  it. 

Thus,  in  blast-furnace  smelting,  it  is  desirable  to  keep  the 
ore  nearly  as  coarse  as  it  comes  from  the-  mine.  (Pieces  as 
large  as  the  double  fist  are  very  suitable  in  the  ore  mixture,  and 
conduce  to  the  openness  of  the  charge  and  the  more  rapid  oper- 
ation of  the  furnace.)  When  in  sampling,  an  aliquot  portion  is 
taken,  as  for  example,  every  fifth  or  every  tenth  shovelful,  the 


OF   THE    COMMON    METALS.  49 

principal  is  broken  no  farther ;  this  is  a  favorite  way  of  sampling 
for  blast-furnace  work.  For  certain  ores  whose  precious-metal 
content  is  high  and  variable,  finer  crushing  is  necessary,  and  the 
blast-furnace  must  make  the  best  of  the  inconvenience  thus  im- 
posed. When  a  suitable  sampling  mill  is  provided,  the  crushing 
of  the  ore  can  be  cheaply  done,  and  the  ore  more  thoroughly 
averaged.  This  crushing  (called  coarse  crushing)  may  be  down 
to  i  in.  or  to  0.75  in.  cubes  and  smaller.  The  sample  taken  from 
this  is  again  crushed  finer,  say  to  o.25-in.  mesh ;  and  after  cutting 
down  the  still  smaller  portion  is  ground  to  say,  3O-mesh.  Finally 
the  very  small  reserved  quantity  intended  to  be  used  as  an  assay 
sample  is  ground  to  pass  an  80  or  loo-mesh  screen. 

Breaking  or  grinding,  to  fit  the  ore  for  further  treatment,  may 
be  divided  into: 

( i ) .  Breaking  for  blast-furnaces. 

(2).  Breaking  for  stall  or  heap  roasting. 

(3).  Crushing  for  reverberatory  melting. 

(4).  Crushing  for  reverberatory  roasting. 

(5).  Crushing  for  leaching  or  percolation. 

(6).  Crushing  for  distillation. 

(7).  Crushing  or  grinding  for  amalgamation. 

Breaking  for  blast-furnaces. 

The  conditions  have  been  referred  to  under  the  head  of  'crush- 
ing for  sampling.'  The  larger  lumps  are  broken  with  6-lb.  ham- 
mers to  pieces  not  exceeding  in  size  the  double  fist.  This  size 
of  hammer,  which  is  more  serviceable  than  a  larger  one,  should 
have  its  faces  rounded  to  increase  the  efficiency  of  the  blow. 
Where  all  the  ore  is  both  oxidized  and  lumpy  the  whole  of  it 
is  passed  through  a  rock  breaker  set  quite  open — say  2^2  in. 
This  is  desirable,  however,  only  when  the  blast-furnace 
charge  is  liable  to  be  too  open.  Quite  commonly  the  reverse 
prevails,  because  of  the  large  amount  of  fine  ore  which  must  be 
run.  .With  a  coarse  charge,  it  is  advisable  also  to  crush  the 
fluxes,  that  is,  the  iron  ore  and  limestone.  When  the  charge  is 
fine,  the  limestone,  at  least,  may  be  fed  in,  in  large  lumps  of  50 
Ib.  and  less.  Limestone  thus  fed  in,  when  heated  in  its  downward 
passage,  finally  disintegrates,  up  to  which  time  it  has  performed 
I'ts  duty  in  making  voids  in  the  charge. 


5O  THE    METALLURGY 

An  objection  to  finely  pulverized  ore  in  the  blast-furnace  is 
that  it  is  likely  to  be  borne  away  by  the  force  of  the  blast.  By 
moistening  the  charge  this  loss  is  lessened ;  but  the  effective  way 
is  to  press  this  fine  material  into  bricks,  which  become  hard  on 
drying,  and  are  unaffected  by  the  blast.  This  adds,  however,  to 
the  expense  of  treatment. 

(2).     Breaking  for  stall  or  heap-roasting,  or  stack-roasting. 

This  work  may  be  performed  either  by  means  of  a  jaw  crusher 
(by  mechanical  means)  or  with  the  spalling  hammer.  By  the 
former  method  the  work  is  done  more  cheaply  but  with  the 
making  of  a  larger  quantity  of  fine  than  in  the  latter  method. 
Peters  cites  a  case*  where,  in  breaking  by  a  jaw  crusher,  the 
fines  amounted  to  17.3%  while  with  hand  hammers  there  was 
only  9%  fines. 

On  the  other  hand,  machine  crushing  cost  9c.  per  ton  while 
by  hand-spalling  the  cost  was  35c.  He  remarks  that,  as  about 
10%  of  the  fine  is  sufficient  for  finishing  the  roast-heaps  or  stalls, 
any  excess  is  to  be  avoided.  Of  course,  this  rule  does  not  apply 
when  the  excess  of  fine  can  be  roasted  in  a  reverberatory  furnace. 

(3).     Crushing  for  rei'erberatory  melting  or  smelting. 

Since  the  ore  is  to  be  fused  into  a  liquid  mass  it  is  not  important 
that  it  should  all  be  crushed  fine.  It  is  well,  however,  to  have 
no  pieces  larger  than  a  hen's  egg.  The  ore  leaves  the  charging 
hopper  more  readily  and  probably  melts  down  more  evenly  than 
in  larger  lumps.  One  great  advantage  of  the  reverberatory  over 
the  blast-furnace  is  that  at  the  time  of  charging,  the  damper  of 
the  furnace  may  be  closed  until  the  dust,  arising  from  dropping 
the  charge,  has  subsided  so  that  there  is  no  serious  loss  of  fine 
or  dusty  ore.  Such  light  material,  in  the  blast-furnace,  would 
be  blown  away. 

(4).     Crushing  for  roasting  in  revcrberatories. 

The  ore  must  be  crushed  fine  enough  to  roast  readily.  An  ore 
containing  iron  sulphide  disintegrates  in  roasting,  so  that. crush- 
ing to  pass  a  2  or  3-mesh  screen  is  sufficient.  Many  ores  and 
mattes  require  finer  crushing,  to  4  and  6-mesh,  for  example.  For 
leaching,  the  ores  are  generally  ground  finer  than  is  necessary 
for  their  efficient  roasting,  though  it  must  be  remembered  that, 

*'Modern  Copper  Smelting,'  by  Dr.  E.  1).  Peters. 


OF    THE    COMMON    METALS.  51 

where  it  is  desired  to  have  a  hot  finish  to  the  roast,  the  smaller 
sizes  roast  more  readily,  and  the  sulphides  are  more  completely 
broken  up.  This  applies  especially  in  the  Stetefeldt  furnace,  where 
the  time  of  roasting  is  limited  to  the  time  occupied  by  the  fall 
of  the  particles  of  ore  to  the  bottom  of  the  tower. 

(5).     Crushing  for  leaching  or  percolation. 

In  this  operation  one  is  governed  by  the  activity  or  efficiency 
of  the  leaching  solution,  whether  of  chlorine,  bromine,  cyanide  of 
potassium,  hyposulphite  of  soda,  salt,  or  simple  water.  With 
an  active  agent  like  chlorine  the  reduction  in  size  may  be  less 
than  is  the  case  with  cyanide  of  potassium.  Again,  the  ore  may 
be  porous,  or  the  precious  metal  may  be  in  a  more  soluble  condi- 
tion or  may  be  rendered  so  by  the  preliminary  roasting.  Thus 
an  ore  crushed  to  lo-mesh  for  chlorination  would  be  broken  to 
3O-mesh  for  cyaniding.  Another  point  to  be  particularly  observed 
in  leaching  is  to  avoid  the  making  of  slimes  in  crushing.  Any 
considerable  quantity  of  this  extremely  fine  ore  hinders  percola- 
tion greatly.  Consequently  gradual  reduction  methods  should  be 
followed  in  crushing  the  ore,  that  is,  the  ore,  already  fine  enough, 
should  be  screened  out  as  fast  as  it  is  formed  and  removed  from 
the  further  action  of  the  crushing  machinery.  On  this  account 
those  machines  where  the  ore  falls  away  from  the  crushing  sur- 
faces are  to  be  preferred. 

(6).  Crushing  for  distillation;  as  in  the  distillation  of  zinc, 
mercury  and  cadmium. 

The  ore  is  broken  down  to  8  or  14-mesh  for  zinc  and  cadmium 
which  is  quite  as  fine  as  is  necessary  for  the  preliminary  roast. 
For  cinnabar  minerals  the  ore  is  fine  enough  when  of  i-in.  size 
or  less. 

(7).     Crushing  for  amalgamation. 

The  ore  is  finely  comminuted  to  pass  through  a  battery  screen 
and  a  good  deal  of  slimed  material  is  produced.  This  is  not  ob- 
jectionable when  the  pulp  passes  over  the  apron  plate.  In  silver 
milling  the  ore  is  finely  ground  in  an  amalgamating  pan,  thus 
comminuting  the  coarser  particles  which  have  got  through  the 
battery  screen.  Grinding  is  continued  until  the  pulp  feels  smooth 
and  free  from  grit  between  the  fingers.  All  of  it  is  practically 
finer  than  2oo-mesh. 


52  THE    METALLURGY 

Dry  crushing. 

In  preparation  of  ore  to  be  finely  ground  for  roasting,  which 
is  to  be  afterward  amalgamated,  or  to  be  treated  by  leaching 
for  the  extraction  of  its  metal,  dry  crushing  is  the  method  fol- 
lowed. 

In  preparation  of  a  considerable  quantity,  say  200  tons  per 
day,  the  following  method  might  be  employed. 

Coarse  crushing. — A  20  by  12  in.  Blake  rock-breaker  will 
crush  25  tons  per  hour  from  12-in.  pieces  dowTn  to  i-5-in.  size 
and  can  take  the  ore  as  it  comes  from  the  mine  at  that  rate. 
However,  there  are  clayey,  talcy,  wet  ores  containing  as  high 
as  25  to  30%  moisture,  which  stick  to  the  machine  and  are  im- 
possible to  crush  in  their  wet  state.  Such  ores  should  be  first 
dried,  using  the  cylinder-dryer  described  below. 

A  set  of  36  by  i6-in.  rolls  will  crush  about  25  tons  per  hour 
through  a  o./5-in.  screen.  This  material  is  now  fine  enough  for 
automatic  sampling,  the  whole  stream  passing  over  a  Yezin 
sampler  (see  Fig.  14)  making  40  rev.  per  min.,  the  rejected 
portion  of  four-fifths  going  to  storage  bins.  The  sample  is  com- 
pleted by  the  well-known  sampling  methods.  One  sampler  ma- 
chine can  handle  200  tons  in  10  hours  of  day-work. 

Fine  crushing. — This  work  is  done  on  both  day  and  night 
shifts,  that  is  to  say  continuously.  The  ore,  0.75  to  o.5o-in.  in  size, 
is  drawn  as  desired,  from  the  storage  bins  by  two-wheeled  buggies, 
and  is  dumped  into  the  hopper  of  the  cylinder-dryer  where  a 
feeding  shoe,  or  other  automatic  feeder,  continuously  supplies  it 
to  the  dryer.  A  cylinder-dryer  24  ft.  long  and  respectively 
4  and  5  ft.  diameter  at  the  two  ends  will  have  sufficient 
capacity  to  take  care  of  ore  containing  an  average  of  6% 
moisture,  at  the  rate  of  10  to  15  tons  per  hour,  drying  it  to  i% 
moisture  or  less,  and  at  the  same  time  heating  it  to  120°  C,  so 
that  it  will  readily  pass  through  the  screen  of  a  trommel.  "When 
in  this  condition  the  ore  is  lively  and  will  screen  readily,  but 
were  it  not  for  this  drying  the  ore  would  stick  to,  and  'blind'  the 
screen.  The  cost  of  such  drying  may  be  put  at  5c.  per  ton. 

For  fine  crushing  a  series-system,  using  four  rolls  (see  Fig. 
21 )  is  adopted  as  follows: 


OF    THE    COMMON    METALS. 


53 


Rolls  a  reduce  the  ore  from  0.75  down  to  0.25  in.  It  is  elevated 
by  a  belt  elevator  to  a  trommel,  having  0.25  in.  perforated  plate 
screens,  the  oversize  going  back  to  rolls  a  while  the  undersize 
passes  on  to  rolls  b.  These  crush  the  material,  which  is  again 


FIG.  21.     FLOW-SHEET  FOR  DRY  CRUSHING. 

elevated  to  a  trommel  with  a  5-mesh  wire  screen.  The  oversize 
of  this  trommel  goes  back  to  rolls  b,  while  the  undersize  is  divded, 
one-half  going  to  each  of  the  two  finishing  rolls  c  and  d.  Here,  as 
before,  the  oversize  of  each  roll,  after  passing  over  a  3O-mesh 


54  THE    METALLURGY 

trommel,  goes  back  to  the  rolls,  while  the  undersize,  being  now 
reduced  to  3<>mesh  (o.O2-in.  particles)  goes  to  the  storage  bin 
m,  and  is  ready  for  further  metallurgical  treatment.  One 
square  foot  of  3O-mesh  screen  will  take  care  of  6  cu.  ft.  of  finished 
product  in  24  hours.  The  desired  quantity  of  200  tons  can  be 
worked  off  in  20  to  22  hours,  thus  allowing  time  for  delays  and 
repairs.  This  gives  the  capacity  of  each  set  of  finishing  rolls  at 
5  tons  per  hour. 

This  system  of  gradual  reduction  is  preferably  followed  be- 
cause the  product  contains  a  minimum  quantity  of  slime,  and 
being  granular,  is  more  easily  percolated  or  leached.  As  each 
piece  or  particle  of  ore  is  crushed  by  a  single  nip,  the  fine  material 
is  screened  out  and  removed  from  further  unnecessary  breaking, 
or  the  consumption  of  more  power  in  so  doing.  In  this  respect 
crushing  by  rolls  has  an  advantage  over  stamps,  in  that  the 
product  is  more  granular  and  uniform.  In  stamp-milling,  how- 
ever, after  the  coarse  crushing  to  1.5  in.  the  rest  of  the  work 
can  be  done  in  a  single  machine  to  the  finest  mesh  desired. 

To  keep  rolls  up  to  their  work  they  should  be  furnished  with 
strong  springs,  with  as  high  as  15  tons  pressure  to  a  bearing; 
that  is  to  say,  the  springs  should  be  such  that  if  a  hard  object 
gets  into  the  rolls,  they  will  only  begin  to  open  when  a  pressure 
of  30  tons  has  been  overcome.  The  speed  of  rolls  varies  with 
the  coarseness  of  the  material  to  be  crushed,  the  coarse  rolls 
running  at  say  600  ft.  peripheral  speed  while  fine  rolls  go  at  i.ooo 
to  1,400  ft.  per  minute. 

Cost  of  dry  sampling  and  crushing  as  above. — The  prime  cost 
of  crushing  ore  in  Colorado  to  3O-mesh,  in  preparation  for  roast- 
ing or  leaching,  may  be  taken  as  follows : 

For  coarse  crushing  and  automatic  sampling.  .  .  .$0.106 

For  fine  crushing,  including  drying °-275 

For  power 0.105 


$0.486 

or  in  round  numbers,  5oc.  per  ton  to  which  must  be  added  cost 
of  management  and  office  expense,  rates  and  taxes,  insurance,  cost 
of  water  supply,  improvement,  and  general  expense. 


OF   THE    COMMON    METALS.  55 

9.     THERMO-CHEMISTRY  As  APPLIED  TO  METALLURGY. 

One  of  the  most  striking  phenomena  of  thermo-chemistry  is 
the  combustion  of  fuel,  in  which  carbonaceous  material  becomes 
united  to  the  oxygen  of  the  air,  the  reaction  being  accompanied 
by  the  development  of  heat.  The  assayer  observes  that  heat  is 
developed  as  the  result  of  cupellation.  The  air  passing  over  the 
surface  of  the  molten  lead  oxidizes  it,  producing  heat  and  enabling 
the  operation  to  progress  at  a  temperature  much  below  the  melt- 
ing point  of  litharge  which  consequently  forms  in  crystals  upon 
the  borders  of  the  cupel.  Again,  a  mixture  of  oxygen  and  hydro- 
gen is  exploded  by  the  momentary  assistance  of  an  electric  spark, 
water  being  formed,  and  the  reaction  being  accompanied  by  the 
development  of  much  heat.  The  heat  generated  in  the  formation 
of  many  chemical  compounds  has  been  determined,  the  unit  by 
which  it  is  measured  being  the  heat  required  to  raise  a  unit  weight 
of  water  one  degree  Centigrade.  The  experimental  determinations 
are  made  with  a  few  grams  of  substance,  and,  hence,  it  has  been 
customary  to  use  as  a  unit  the  heat  required  to  raise  the  tem- 
perature of  one  gram  of  water  to  one  degree  Centigrade.  This 
unit  is  called  a  small  calorie  (cal).  The  amount  of  heat  thus 
taken  is,  however,  too  small  for  practical  application,  so  that  it 
has  been  customary  to  use  the  heat  needed  to  raise  one  kilogram 
of  water  to  one  degree  Centigrade,  called  a  large  calorie  (Cal), 
or  one  thousand  times  the  other  quantity.  Where  the  unit  weight 
is  one  pound,  we  have  the  pound-calorie,  in  which  that  amount 
of  water  is  also  raised  one  degree  Centigrade.  This  unit  will 
generally  be  used  in  the  calculations  to  follow.  When  12  Ib.  of 
carbon  (C)  are  burned  with  32  Ib.  of  oxygen  (2O)  there  results 
44  Ib.  of  carbon  dioxide  (CO.,),  and  this  is  expressed  by  the 
chemical  equation  (i)  C-J-2O=CO2,  that  is  to  say,  97,000  calories 
are  developed,  or  8,080  calories  for  each  pound  of  carbon  burned. 
It  must  be  understood,  in  writing  equations,  that  the  molecular 
weight  is  referred  to  as  expressed  in  an  ordinary  chemical  equa- 
tion. This  may  also  be  indicated  thus:  (2)  C,  O.2  =  97,000, 
meaning  that  the  separate  molecules  C  and  O  are  thus  united. 
If  we  burn  carbon  with  a  limited  quantity  of  oxygen,  as  in  the 


56  THE    METALLURGY 

upper  part  of  a  thick  fire,  CO  gas  results  thus  (3)  C-f-O=CO, 

29,000 

or,  as  otherwise  expressed,  (4)  C,  0=29,000,  or  the  heat  devel- 
oped by  the  incomplete  combustion  of  carbon,  is  but  2440  Ib. 
calories.  If  we  burn  the  CO  gas,  thus  formed,  with  an  excess 
of  air  we  again  obtain  carbon  dioxide,  thus  (5)  CO-|-O=CO2. 

68,000 

This  means  that  before  the  new  compound  can  be  formed  the 
CO  must  be  decomposed,  using  up  in  so  doing  the  heat  just 
obtained.  The  equation  may  also  be  given  (6)  CO,  O=68,ooo 
the  comma  showing  what  are  the  elements  to  be  combined.  It  will 
be  noted  that  the  heat  developed  by  the  burning  of  the  carbon 
in  two  stages,  viz:  (6)  29,000+68,000=97,000  amounts  to  the 
same  as  though  the  carbon  had  been  at  once  completely  burned 
to  carbon  dioxide.  We  thus  get  the  following  laws : 

1.  The  amount  of  heat  needed  to  decompose  a  compound  into 
its  constituents  is  equal  to  that  evolved  when  that  compound  is 
formed  from  those  constituents.     When  a  reaction  takes  place 
by  which  heat  is  absorbed,  as  in  the  decomposition  just  referred 
to,  it  is  called  endothermic.     On  the  other  hand  where  heat  is 
evolved  the  reaction  is  an  exothermic  one. 

2.  The  heat  evolved  in  a  chemical  process  is  the  same  -whether 
it  takes  place  by  one  or  b\  several  steps.     See  equations  (5)  and 
(6). 

In  comparing  reactions  (i)  and  (3)  it  may  be  said  that  in  pres- 
ence of  an  excess  of  oxygen,  the  reaction  (i),  by  which  CO2  is 
formed,  will  take  place  rather  than  reaction  (2)  in  accordance 
with  the  law  of  Berthelot,  namely: 

3.  Every  reaction  'which  takes  plate  independently  of  the  ad- 
dition of  energy  from  without  the  system,  tends  to  form  the 
combination  which  is  accompanied  by  the  greatest  evolution  of 
heat. 

The  reaction  produced  by  the  oxidation  of  lead  to  litharge,  re- 
ferred to  above  is  thus  expressed :  Pb  +  O,  =  PbO  or  otherwise 
Pb,  O  =  51,000;  but  on  account  of  its  high  atomic  weight  this  is 
but  249  Cal.  per  pound  of  lead. 

To  determine  accurately  the  heats  of  formation  of  fuels  and 
other  compounds,  the  Mahler  bomb  calorimeter  (Fig.  22)  is  much 


OF    THE    COMMON    METALS. 


57 


used.  It  consists  of  a  steel  shell  or  bomb  B  having  a  capacity  of 
over  a  pint.  The  shell  is  closed  by  a  screw-cap,  with  a  connection 
X  for  the  introduction  of  oxygen  from  a  cylinder  of  compressed 
oxygen  gas  O.  Within  the  bomb  is  suspended  a  capsule  C,  in 
which  is  placed  the  substance  to  be  tested.  The  bomb  is  then 
closed,  placed  in  the  water  of  the  calorimeter  A,  and  charged  with 
the  compressed  oxygen.  The  fuel  is  ignited  by  the  aid  of  a  wire 
F,  which  is  heated  to  redness  by  an  electric  current.  An  ex- 
plosion takes  place,  and  the  heat  of  the  completely  burned  fuel 
is  transmitted  to  the  water  of  the  calorimeter.  The  water  sur- 
rounding the  bomb  is  kept  in  circulation  by  means  of  a  mixer  S, 


FIG.  22.     MAHLER  BOMB  CALORIMETER. 

and  the  rise  of  temperature  noted  by  the  thermometer  T.  The 
total  heat  developed  is  calculated  with  certain  corrections  from 
the  amount  of  calorimeter  water  and  the  rise  of  temperature. 
In  those  cases,  where  the  formation  temperatures  of  oxides  or 
silicates  are  desired,  it  is  accomplished  by  melting  or  oxidizing 
the  compound  in  the  bomb  with  a  known  weight  of  a  well-de- 
termined fuel.  The  total  calories  evolved  would  be  the  algebraic 
sum  of  those  of  the  desired  reaction  and  of  the  fuel. 

Heat  of  formation. — The  amount  of  heat  liberated  or  absorbed 
in  the  formation  of  a  molecule  of  a  substance  is  called  the  heat  of 


58  THE    METALLURGY 

formation,  and  is  expressed,  either  in  the  large  pound-calorie,  or 
in  the  small  gram-calorie.  If  the  heat  of  combustion  of  a  com- 
pound be  subtracted  from  the  heats  of  combustion  of  its  elements, 
the  remainder  is  its  heat  of  formation.  The  heat  of  forma- 
tion of  methane  (CH4)  is  determined  indirectly  from  the  heat  of 
combustion  thus : 

CH4+4  O=CO2-f2  H2O=i9i,ooo 
97,000     2X58,000 

The  heat  of  formation  of  methane  is  therefore  (97,ooo-(-2X 
58.000) — 191,000=22,000  calories. 

While  the  equations  heretofore  given  are  exothermic,  it  may 
often  happen  that  a  reaction  takes  place  accompanied  with  the  ab- 
sorption of  heat,  as  in  the  reduction  of  metals.  Thus,  we  have 
a  reaction  which  takes  place  only  at  a  red  heat  because  of  the 
large  amount  of  heat  absorbed  in  the  reaction : 

2  Pb  SO4+2  PbS=2  PbO+Pb+PbS+3  SO2 

2X216200  2X17800  2X51000  17800  3X71000= — 135200 


PART  II.    ROASTING 


PART    II.     ROASTING. 


10.    THE  CHEMISTRY  OF  ROASTING. 

For  purposes  of  illustration  let  us  consider  the  reverberatory 
i  casting  of  an  ore  containing  mixed  sulphides,  galena,  blende, 
pyrite  and  chalcopyrite,  together  with  a  quartz  gangue,  which 
prevents  cohesion  of  the  particles  of  ore  and  makes  the  charge 
more  open  and  accessible  to  air.  The  ore  is  first  dropped  upon 
the  hearth  at  the  cooler  end  of  the  furnace  where  the  temperature 
(35o°C.)  is  only  sufficient  to  drive  off  moisture  and  to  start  the 
reactions.  In  10  to  15  minutes  the  charge  becomes  hot  enough 
for  oxidation  of  the  pyrites  to  take  place,  as  shown  by  a  blue 
flickering  flame  which  plays  upon  the  surface  of  the  charge.  The 
second  equivalent  of  sulphur  in  FeS2  is  feebly  held,  so  that  it 
comes  off,  burning  to  SO2  according  to  the  reaction, 
(i).  S+2O=SO2 
+71000 

or    71000  -f-  32  =  2220    pound-calories    per    pound    of    sulphur 
burned. 

At  the  same  time,  the  FeS  remaining,  together  with  the  other 
sulphides  CuS,  ZnS  and  PbS,  begins  to  oxidize,  especially  the 
FeS,  which  acts  according  to  the  reaction : 

(2)  FeS  +  30  =  FeO  +  SO2 

23800  66400   7 1 000= +  11 3600 

or,  per  pound  of  sulphur,  113600 

—  —  3550  pound-calories, 
o 

The  cupric  sulphide  of  the  chalcopyrite  acts  according  to  the 
formula : 

(3)  CuS  +  30  =  CuO  +  S02 

10200  37200    7 1 000= -(-98000 

or,  per  pound  of  sulphur,  98000 

—  =  3030  calories. 


62  THE    METALLURGY 

The  blende,  under  the  action  of  the  air,  is  affected  in  the 
same  way: 

(4)  ZnS+3O=ZnOH-SO2 

43000        86400  / 1 000= -(-109400 
or,  per  pound  of  sulphur,  3420  calories. 

Galena  also  roasts  according  to  the  formula : 

(5)  PbS+30=PbO+S02 

17800  51000  7iooo=  +  iO42oo 

from  which  we  obtain  3250  calories  per  pound  of  sulphur.  It 
will  be  noticed  that  the  heat  evolved  per  pound  of  sulphur  is 
much  the  same  in  each  case;  and  hence  such  sulphides  as  FeS, 
containing  much  sulphur,  give  off  the  most  heat  in  roasting.  The 
above  reactions  act  superficially  and  their  activity  is  increased  by 
stirring  the  charge.  Naturally,  FeS  is  first  formed  where  most 
exposed  to  the  air,  which,  reacting  upon  it,  forms  iron  sulphate. 

(6).    3  FeS+n  O=2  SO2-f  Fe2O3+FeSO4 

3X23800      2X71000      199400      23  5600=4-  5°56oo 
or,  per  pound  of  sulphur  5260  cal.,  a  most  energetic  exothermic 
reaction. 

As  the  ore  is  moved  to  a  hotter  portion  of  the  furnace  the 
activity  of  these  reactions  continues,  keeping  up  the  temperature ; 
but  at  the  just  visible  red  (59O°C)  the  just  formed  iron  sulphate 
begins  to  decompose,  reacting  on  the  cupric  oxide : 

(7)  FeSO4+CuO=FeO+CuSO4 

235600  37200  66400     182600=: — 23800 
and 

(8)  FeSO4+2  CuO=FeO+CuOCuSO4 
2356002X37200   66400     145400=— 98200 

There  remains  unused,  however,  much  iron  sulphate  which  is 
decomposed  at  the  higher  temperature,  thus : 

(9)     FeSO4=FeO+SO3 

235600   66400  91800=— 77400 

It  will  be  noticed  that  these  latter  three  reactions  (7),  (8),  and 
(9)  are  endothermic.  At  a  slightly  greater  heat  (655°  C.)  the 
just  formed  cupric  sulphate  (reaction  (7))  begins  to  decompose; 
and  at  a  dark  red  heat  (705°  C.)  the  same  change  begins  upon  the 


OF    THE    COMMON    METALS.  63 

cupro-cupric  sulphate  (reaction  (8) ).  These  decompositions  of  the 
copper  sulphates  are  complete  at  850°  C.  or  at  a  cherry-red  heat. 
At  this  time  also  the  zinc  and  lead  oxides,  reacting  with  the  copper 
sulphates,  are  changed  to  sulphates : 

ZnO+CuSO4=ZnSO4+CuO 

86400  182600  230000  37200= — 1800 

PbO+CuSO4=PbSO4-fCuO 

51000   182600    216200     37200= — 19800 

As  the  heat  increases  these  sulphates  again  decompose,  the  zinc 
sulphate  more  readily  than  the  lead  sulphate.  At  i,o5O°C,  or  a 
dark-orange  heat,  copper  oxide  is  reduced  to  cuprous  oxide 
(Cu0O)  and  ferric  oxide  to  the  next  lower  oxide,  Fe3O4. 

At  this  stage  the  ore  begins  to  fuse  together,  or  agglomerate 
slightly,  and  is  then  withdrawn  from  the  furnace.  If  much  lead 
exists  in  the  ore,  even  this  temperature  cannot  be  attained  without 
the  ore  beginning  to  soften,  when  roasting  must  cease  because 
the  charge  is  no  longer  porous.  In  such  case  the  lead  and  zinc 
sulphate  are  imperfectly  decomposed,  and  the  sulphur  is  not  well 
eliminated. 

Sometimes,  and  with  zinciferous  ores,  the  roaster  is  arranged 
with  a  fuse-box,  into  which  the  charge  is  brought  and  melted 
down  with  the  addition  of  some  silicious  ore,  the  silica  reacting 
on  the  zinc  and  lead  sulphates  according  to  this  formula: 

ZnSO4+SiO2=ZnSiO3+SO3,  and 
PbSO4+Si02=PbSiO3+SO3 

the  sulphur  being  eliminated  as  sulphuric  anhydride. 

It  has  been  found  that  with  2%  by  volume,  or  4.4%  by  weight 
of  SO2  in  the  escaping  gases,  roasting  proceeds  actively.  This 
corresponds  to  23  Ib.  air  per  pound  of  SO2,  or  to  46  Ib.  of  air  per 
pound  of  sulphur  driven  off.  If  more  air  is  admitted  than  this, 
then  the  charge  gets  cooled ;  if  less,  roasting  proceeds  slowly ; 
with  8%  SO2,  it  ceases  altogether.  Thus,  not  only  must  the  heat 
be  sufficient,  and  new  surfaces  presented  to  the  action  of  the  air, 
but  the  air  must  be  fairly  fresh  and  the  SO2  evolved  must  be 
promptly  removed  from  contact  with  the  ore  by  means  of  the 
draft. 


64  THE    METALLURGY 

The  various  reactions  described  above  need  time  for  their  com- 
pletion ;  and  the  larger  the  body  of  ore,  the  longer  it  takes  to 
complete  the  roast.  If  a  few  grams  of  ore  are  roasted  in  the 
muffle  the  operation  is  complete  in  half  an  hour,  but,  in  a  rever- 
beratory  furnace  roasting  14  tons  per  day  the  same  operation 
takes  twenty  hours. 

The  temperatures  at  which  the  various  reactions  take  place  are 
as  follows : 

At  350° C.  the  sulphur  of  sulphides  begins  to  burn  off. 

At  59O°C.  just  formed  iron  sulphate  (FeSO4)  begins  to  de- 
compose. 

At  655 °C.  just  formed  copper  sulphate  (CuSO4)  begins  to  de- 
compose. 

At  705  °C.  cupro-copper  sulphate  (CuOCuSO4),  formed  at  the 
same  time  as  the  copper  sulphate  (Cu  SO4),  begins  to  decompose. 

At  850° C.  copper  sulphates  are  quite  decomposed. 

At  835°  to  850°  C.  the  maximum  amount  of  soluble  silver  sul- 
phate (AgSO4)  is  formed. 

At  1050° C.  copper  oxide  (CuO)  is  reduced  to  Cu2O. 

At  uoo°C.  ferric  oxide  Fe2O3  is  reduced  to  the  next  lower 
oxide  Fe3O4. 

Losses  in  roasting. — This  depends  upon  the  extent  to  which 
the  roast  is  carried  as  well  as  the  nature  of  the  ore.  If  the  ore 
is  dry-roasted,  so  that  it  is  not  sintered  at  all,  the  lead-loss  will 
be  2.5%,  with  no  loss  in  silver.  When  the  ore  is  agglomerated 
the  losses  are  slightly  higher;  and  when  fused  the  losses  are  15 
to  20%  Pb,  2  to  5%  Ag. 

The  lead-bearing  matte  from  the  blast-furnace  is  also  roasted, 
preferably  in  a  reverberatory  roaster.  It  requires  a  different 
treatment  from  ore  since  it  contains  but  20%  S,  does  not  take 
fire  like  a  pyrite  ore,  and  must  have  a  good  finishing  heat  in 
order  to  expel  sulphur.  Matte  is  considered  to  be  well  roasted 
when  still  containing  4%  S.  Ores  low  Tn  lead  may  be  easily 
brought  down  to  2  to  3%  sulphur,  while  galena  still  retains  5 
to  6%  S  when  withdrawn  from  the  furnace.  Like  matte,  galena 
starts  burning  slowly,  and  must  be  slowly  roasted — premature 
heating  at  once  sintering  it,  thus  stopping  further  roasting. 


OF    THE    COMMON    METALS.  65 

The  following  is  an  analysis  of  a  leady  matte  (unroasted)   from 
Pueblo,  Colorado: 

10.72%   PbO  0.56%       As 

0.61       Cn  0.41       CaO 

24.01       S  0.47       MgO 

52.27       Fe  24.4  oz.  Ag  per  ton 

4.27       Zn 

with  a  ratio  of  sulphur  to  iron  of  I  to  2.2. 

ii.    ROASTING  OF  ORES  IN  LUMP  FORM. 

Heap  roasting. — The  roasting  in  heaps  of  sulphide  ores  con- 
taining copper,  is  an  operation  that  must  be  performed  with 
knowledge  and  care  in  order  to  obtain  satisfactory  results.  The 
selection  of  this  method  involves  a  careful  consideration  of  the 
environment.  Thus,  it  may  not  be  adopted  in  a  settled  country, 
where  the  fumes  from  it  are  a  nuisance,  or  where  it  is  likely  to 
damage  cultivated  crops,  surrounding  vegetation,  or  live-stock. 
In  the  arid  and  sparsely  settled  regions  of  the  Rocky  Mountain 
States,  it  may  often  be  used  to  advantage.  At  a  moderate-sized 
installation,  where  not  more  than  25  tons  of  sulphur  are  evolved 
daily,  the  zone  affected  may  not  be  more  than  4  miles  in  extent, 
and  the  roasting  site  may  then  be  chosen  with  this  in  view. 
Often  prevailing  winds,  blowing  from  a  single  quarter,  may  per- 
mit the  placing  of  roast-piles  so  as  to  give  practically  no  offense. 
Indeed,  this  matter  should  be  considered  with  relation  to  the 
plant  itself  and  the  ground  chosen,  so  that  the  smoke  shall  be  sel- 
dom driven  in  the  wrong  direction.  A  roast-yard  should  be  of  ample 
size,  approximately  level,  and  protected  from  flow  of  storm- 
water.  At  Jerome,  Arizona,  a  tram  track  following  the  contour 
of  the  hillside  has  the  roast  heaps  adjoining  it,  so  that  the  ore 
is  conveniently  transferred  to  the  ground,  and  the  roasted  ore 
carried  to  the  smelter  by  the  same  tram-system.  This  tends  also 
to  scatter  the  heaps,  so  that  fumes  do  not  interfere  with  making 
them  up. 

An  ordinary  pile,  say  40  ft.  long,  24  ft.  wide  and  6  ft.  high, 
and  containing  240  tons,  will  burn  for  70  days,  to  which  is  to 


66  THE    METALLURGY 

be  added  10  days  for  removing  and  rebuilding.  This  will  make  a 
yield  of  3  tons  of  roasted  ore  daily.  A  foundation  should  be 
prepared  by  leveling  off  the  ground,  and  making  a  final  surface  of 
clayey  loam.  Upon  this  is  placed  a  layer  of  ore-fine  to  the  depth 
of  3  or  4  in.  As  this  layer  gets  gradually  roasted,  it  is  sent  to 
the  furnaces,  fresh  fine  taking  its  place.  A  system  of  overhead 
tram-trestles  (see  Fig.  23  and  24)  may  often  be  put  in,  other- 
wise the  ore  is  hauled  by  carts  or  brought  in  cars  to  the  heap, 
and  is  then  wheeled  on. 

The  bents,  shown  in  the  figure,  are  36  ft.  apart,  and  trussed 
stringers,  10  by  12  in.,  carry  the  tram  track  over  the  roast-heaps. 
A  turn-plate,  with  a  movable  track,  serves  to  carry  the  ore  to  the 
full  length  of  the  roast-heap.  This  track,  made  with  stout  rails, 
is  supported  on  movable  trestles.  The  height  of  the  pile  will 
depend  upon  the  character  of  the  ore.  Thus,  an  ore  of  15% 
sulphur  may  be  made  9  ft.  high,  while  massive  pyrite  should  be 
but  6  ft.  in  height  for  the  best  conditions  of  thorough  roasting. 

Upon  a  bottom  of  the  fine  is  placed  the  fuel  used  to  start  the 
roasting.  This  consists  of  a  layer  of  wood  of  4  to  8  in.  in  thick- 
ness, the  4-in.  layer  being  sufficient  for  the  higher  sulphide  ores. 
The  wood  may  be  of  any  kind  and  length,  especially  when  these 
higher  sulphides  are  roasted.  The  cheaper  wood,  as  old  rails, 
logs,  old  tree  trunks  and  twisted  branches,  can  be  taken  for  the 
central  portion  of  the  pile,  using  the  more  uniform  wood  for 
the  outer  four  feet  of  the  borders.  The  interstices  between  these 
closely  packed  pieces  should  be  closed  by  laying  in  finer  sticks, 
brushwood  and  chips,  so  that  the  ore  will  not  fall  through.  Three 
chimneys,  8  in.  square,  made  of  four  old  boards  set  upright  from 
the  ground,  are  located  evenly  on  the  foundation,  with  channels, 
say  6  in.  wide,  containing  kindling  wood  connecting  the  borders 
to  the  chimneys.  The  chimneys  may  also  be  made  of  sticks  wired 
together,  or  old  sheet  iron  bent  in  cylindrical  form.  The  coarse 
ore,  spalled  so  that  none  of  the  pieces  are  more  than  4  in.  in 
size  and  down  to  I  in.  diameter,  is  dumped  upon  the  foundation. 
The  ore  is  carried  to  the  required  height  with  as  steep  an  angle 
as  it  will  stand,  forming  a  shapely  frustum  of  a  pyramid  with 
sharp  corners.  Upon  this,  as  shown  in  Fig.  23,  is  placed  the 
ragging  (from  i  in.  down  to  0.25  in.  in  size),  forming  a  layer 


OF   THE    COMMON    METALS.  67 

thicker  below  and  thinning  out  as  we  go  up  the  pile.     Following 
this,  a  thin  layer  of  fine  is  added  on  the  slopes  of  the  pile. 

It  should  be  noted  here,  that  far  too  much  wood  is  apt  to  be 
used.  About  one  cord,  on  an  average,  is  required  for  40  tons  of 
ore.  It  must  be  remembered  that  the  object  of  the  wood  is 
merely  to  start  the  pile  burning.  As  the  wood  soon  burns  out, 
the  pile  will  continue  to  burn  by  its  own  action.  The  more  mas- 
sive the  pyrite  the  less  wood  required.  While  the  pile  should  be 


FIG.  23.     ROAST-YARD  WITH   TRESTLE    (CROSS   SECTION). 


FIG.  24.     ROAST-YARD  WITH  TRESTLE  (LONGITUDINAL  SECTION). 

kept  uniformly  burning,  no  attempt  should  be  made  to  unduly 
hasten  the  process,  since  with  a  high  heat  the  ore  will  fuse  to- 
gether, thus  stopping  all  further  action. 

The  pile  is  fired  at  the  different  channels  all  around  the  edge, 
and  preferably  in  fine,  still  weather.  After  4  to  6  hours,  the 
fire  having  spread  over  the  entire  area,  the  fine  is  put  upon  the 
pile,  thinly  on  top  and  more  thickly  upon  the  sides  and  upon  the 
borders  of  it  below.  The  heap,  especially  at  first,  must  be  closely 
watched, — fine  being  placed  to  check  the  draft  where  it  is  too 
vigorous,  and  holes  opened  in  the  covering  to  draw  the  fire  wher- 
ever the  draft  seems  dead.  The  fine  is  freely  used  to  control 
the  fire,  both  on  the  borders  and  the  top  of  the  heap.  The  latter 
can  be  reached  for  examination  and  adjustment  by  taking  ad- 
vantage of  the  wind  and  the  drafts  of  air. 


68  THE    METALLURGY 

Moderate  rains  and  snow  have  but  little  effect  on  the  process, 
but  high  wind  from  one  direction  is  apt  to  stop  burning  on  the 
windward  side.  This  may,  however,  be  prevented  by  a  temporary 
fence  placed  as  a  wind-break.  More  abundant  rains  tend  to  leach 
out  copper  sulphate  from  the  roast  pile.  Where  such  conditions 
exist,  the  heaps  should  be  roofed  over,  or,  where  this  is  thought 
to  be  too  great  an  expense,  some  saving  can  be  effected  by  drain- 
ing through  ditches  to  a  launder  containing  scrap  iron,  where 
the  copper  can  be  precipitated.  In  the  dry  region  of  the  West, 
these  considerations  are  less  important.  In  places  like  Mexico, 
probably  it  would  be  best  to  suspend  operations  during  the  rainy 
season. 

Wherever  possible,  roast  heaps  should  be  left  undisturbed ;  but 
if  ore  must  be  used,  it  may  be  taken  from  the  burnt-out  and 
cooled  portion  of  the  pile,  leaving  action  to  continue  on  the  hot 
core,  still  burning.  The  best  way,  however,  is  to  start  roasting 
operations  weeks  in  advance  of  the  smelting  so  as  to  have  a  well 
roasted  supply  to  draw  from.  With  care  and  experience  it  should 
be  possible  to  roast  90%  of  the  ore  placed  on  the  pile,  including  the 
fine. 

Cost  of  heap  roasting.  We  give  the  cost  of  roasting  at  Duck- 
town,  Tenn.,  at  42c.  per  ton,  while  Peters  gives  48.5^  per  ton 
for  the  items  of  fuel,  labor,  and  supplies.  In  the  latter  case  com- 
mon labor  was  estimated  at  $1.50  per  day.  Heap  roasting  can 
often  be  contracted  for  to  advantage.  At  the  United  Verde  at 
Jerome,  Arizona,  75c.  per  ton  was  paid  on  contract.  Heap  roast- 
ing, skillfully  conducted,  will  reduce  the  sulphur  from  40%  and 
over  to  as  low  as  7  and  8%.  The  roasted  product  has  an  earthy, 
irregular  surface  of  a  blackish  brown  hue,  and  is  lighter  to  the 
hand  and  more  porous  than  the  raw  ore. 

Heap  roasting  of  matte.  Matte  in  lump  form  can  be  well 
roasted  in  heaps,  but,  unlike  ore,  requires  two  or  more  burnings. 
After  the  first  firing,  in  spite  of  the  greatest  care,  the  matte  shows 
but  little  of  the  profound  change  it  has  undergone.  At  the  second 
burning,  with  a  somewhat  larger  quantity  of  wood,  the  results 
begin  to  show,  a  large  portion  of  the  twice-burned  material  then 
appearing  light,  porous  and  with  no  raw  interior.  In  fact,  the 
thoroughness  of  the  roast  may  be  determined  by  the  feeling  of 


OF    THE    COMMON    METALS.  69 

the  lumps  in  the  hand.  If  such  lumps  are  broken  open,  they  no 
longer  have  a  raw  core  or  centre. 

The  wood  bed  for  matte  can  be  prepared  as  for  ore,  but  the 
pile  is  only  12  ft.  square  by  6  ft.  deep,  with  a  single  chimney  in  the 
centre.  The  broken  matte  with  its  fine  constitutes  the  heap, 
which  is  covered  by  thoroughly  roasted  ore  which  is  fine  enough 
for  the  purpose.  The  burning  of  the  heap  will  take  10  days,  after 
which  the  heap  is  broken  up,  and  the  partly  roasted  material  made 
into  a  new  pile.  It  is  a  good  plan,  in  this  new  pile,  to  introduce 
one  or  two  layers  consisting  of  chips  and  bark,  which  have  a 
reducing  effect  on  impurities,  and  also  furnish  more  uniform 
heat  to  the  mass.  Finally  at  the  next  turning  over,  a  large  portion 
suitable  for  use  may  be  sorted  out,  the  partially  burned  remain- 
der going  to  a  fresh  heap.  Peters  gives  the  cost  per  ton  for 
roasting  matte  as  $2.05  for  the  three  burnings. 

Relative  advantages  of  heap  and  of  stall  roasting.  Heap  roast- 
ing has  the  advantage  that  it  requires  only  the  necessary  site  and 
no  investment  for  plant.  The  method  is  a  simple  one  and  the 
results  satisfactory.  On  a  small  scale  primitive  methods  of 
handling  materials  are  sufficient,  but  on  a  larger  scale  we  must 
not  forget  the  costs  of  grading,  for  trestles,  trackage,  etc.  Stall 
roasting  saves  much  time,  requiring  10  days,  as  against  70  days 
for  heap  roasting.  But  in  large  plants,  where  from  10,000  to 
50,000  tons  are  in  process  of  treatment,  several  hundred  thousand 
dollars  may  be  locked  up  in  the  heaps.  By  reducing  this  value  to 
one-seventh  an  important  saving  is  effected.  In  stall  roasting 
the  stack  removes  the  fumes,  and  the  entire  contents  of  the  stall 
get  roasted.  Rain  and  snow  have  but  little  effect  on  the  roasting 
ore,  and  in  a  moist  climate,  leaching  does  not  cause  trouble.  For 
stall  roasting  one-fifth  of  a  cord  of  wood  is  enough  for  the  charg- 
ing of  a  stall,  or  one  per  cent  of  the  ore  to  be  roasted,  while,  in 
heap  roasting,  average  practice  calls  for  2y2%  of  wood. 

Cost  of  roasting  stalls.  Peters  gives,  for  the  cost  of  building 
56  stalls,  a  total  of  $3,303.80,  or  about  $60  per  stall.  The  total 
includes  $448.80  for  the  necessary  trackage.  He  estimates  the 
cost  of  the  track  laid  with  12-lb.  rails  at  5ic.  per  foot.  To  this 
should  be  added  about  $400  as  the  cost  of  the  stack. 

Cost  of  roasting  in  stalls.     This  may  be  assumed  at  5oc.  per 


7<3  THE    METALLURGY 

ton,  using  the  same  figures  for  labor  as  in  calculations  for  heap 
roasting. 

Stall  roasting.  The  principle  of  such  roasting  consists  in  en- 
closing the  ore  within  walls  or  rooms  called  stalls,  in  the  walls  of 
which  are  contrived  flues  for  the  admission  of  air,  and  for  the 
removal  of  smoke,  by  means  of  a  main  flue  to  a  tall  stack,  thus 
shortening  the  period  of  burning,  and  at  the  same  time  removing 
the  noxious  fumes  to  the  upper  air.  The  stall  consists  of  a  small 
paved  area  surrounded  by  three  walls,  but  having  an  open  front, 
which,  to  confine  the  contents,  is  loosely  built  up  at  each  operation. 
The  back,  or  sides,  are  pierced  with  small  openings  communicating 
to  the  common  flue  of  manv  such  stalls.  The  surface  of  the  ore 


CROSS  SECTION  THROUGH  A.B. 
SCALE  X  IN.  =  1  FOOT 

i  i  i  i  i  ,  JL  .  &  ,  fc- 

"T 

1 

b 

si 
J 

^7^5      I/ 
\b.  

i  a.-Dravght  holes  connecting 
with  Flue  in  sideioalls. 
b.b.—Flve  holes  into  Main 
Culvert. 

"a     a      1      1      1 

TRACK  TO 
SMELfER 
_»              * 

i      i     A     i      i      :     f, 

i      i  D  i      i      i  pr_-_:~ 
A    ^     i      i      i      f 
Plcpi      i      i      |  L_?__j 

,|| 

.    ,  1    ,  , 
»    \      \ 

SE 

i 
a 

\ 

TRACK  TO 
SMELTER 

'  &  k-2#—  , 

^4-*^ 

> 

FIG.  25.    ROASTING  STALLS  FOR  LUMP  ORE  (SECTIONAL  ELEVATION). 

in  the  filled  stall  is  covered  with  fine  to  keep  in  the  heat,  and  to 
cause  a  suitable  draft  from  below,  through  the  flues,  to  the  stack. 
Openings  near  the  bottom  permit  the  entrance  of  the  air  which 
also  filters  through  the  front  wall. 
Roasting  Stalls  for  Lump  Ore. 

Fig.  25  and  26  represent,  in  sectional  elevation  and  in  plan,  a 
battery  of  stalls  for  roasling  ore.  Fifty-six  such  stalls  will  roast 
loo  tons  of  raw  ore  daily,  each  holding  20  tons,  taking  10  days  to 
burn  and  clear  out,  or  2  tons  daily  per  stall  with  a  12%  allowance 
of  time  for  repairs.  The  stall  may  be  built  of  rough  masonry  laid 
in  clay  mortar  or  with  slag-blocks  cast  at  the  works  itself.  The 
brick  stack,  3  ft.  6  in.  square  inside,  is  75  ft.  high. 


OF   THE    COMMON    METALS. 


72  THE    METALLURGY 

To  fill  a  stall,  a  central  longitudinal  and  two  lateral  flues  or 
passages  for  the  admission  of  air  are  made  with  large  irregular 
pieces  of  ore.  These  are  filled  and  surrounded  with  kindling 
wood  while  the  rest  of  the  bottom  is  covered  in  a  thin  layer  with 
long  thin  sticks,  split  from  old  logs  and  poles.  The  structure  is 
now  filled  with  coarse  ore,  the  ragging  being  distributed  through 
the  mass.  As  the  stall  is  filled,  single  small  sticks  of  wood  are 
put  at  the  back  and  side-walls  as  well  as  occasional  sticks  at  the 
front.  A  single  carload  of  ragging  is  now  added  above  the  ore, 
then  a  3-in.  layer  of  shavings,  bark  and  chips,  a  layer  of  1.5  tons 
of  fine,  and  finally,  a  coating  of  well-roasted  ore.  Sometimes  a 
sheet-iron  cover  luted  with  clay  at  the  walls,  on  the  top  of  the  ore, 
is  of  great  advantage.  By  care  it  is  possible  to  roast  a  propor- 
tion of  fines  in  this  way,  thus  disposing  of  a  product  which  would 
otherwise  have  to  be  roasted  in  a  reverberatory  furnace.  The 
stall  having  been  fired,  the  roasting  proceeds  rapidly,  and  by 
the  end  of  the  fourth  day  the  heap  should  be  burning  throughout. 
If  the  process  is  successful,  it  will  be  indicated  by  the  swelling  or 
rising  of  the  contents  of  the  stall,  sometimes  to  the  extent  of 
12  in.  Because  of  the  swelling,  the  front  or  temporary  wrall 
should  also  be  braced  to  oppose  the  outward  thrust.  When  the 
process  goes  on  too  rapidly,  no  such  swelling  occurs,  but  on  the 
contrary  the  surface  subsides  where  the  ore,  owing  to  the  great 
heat,  has  melted  together,  and  the  roasting  is  imperfect.  The 
heat  can,  however,  be  regulated  by  the  use  of  fine  for  stopping 
cracks,  and  by  the  closing  of  draft-openings.  Were  the  ore  left 
to  burn  out  and  cool  at  leisure,  it  would  take  15  days.  In  order 
to  hasten  the  operation  the  front  portion  of  the  ore  may  be 
removed  as  it  cools  from  the  front,  taking  care  not  to  penetrate 
beyond  the  cooled  portion.  Beginning  at  about  the  fourth  day, 
it  is  possible  to  take  away  the  ore,  so  that  in  5  to  7  days  the  stall 
is  again  empty.  As  regards  sulphur,  stall  roasting  is  perhaps  a 
little  less  efficient  than  heap  roasting. 

12.    ROASTING  OF  ORES  IN  PULVERIZED  CONDITION. 

This  work  is  performed  in  furnaces,  the  ore  being  exposed  upon 
a  hearth  to  the  action  of  the  hot  gases  and  air  from  the  fire.  We 
mav  divide  them  into: 


OF    THE    COMMON    METALS.  73 

(i).  Hand-rei'erberatory  roasting  furnaces  or  calciners  (in- 
termittent discharge). 

(2).  Revolving  cylinders  with  (a)  continuous  discharge,  as 
the  White-Howell  and  the  Argall;  (b)  intermittent  discharge, 
as  the  Bruckner. 

(3).  Automatic  reverberatory  roasters  or  calciners  with  con- 
tinuous discharge  having  (a)  straight  hearths,  as  the  Brown- 
O'Hara  and  the  Wethey;  (b)  curved  or  circular  hearths,  as 
the  Brown  Horseshoe,  the  Pearce  Turret,  the  McDougall  and  the 
Herreshoff. 

Besides  these  should  be  mentioned  the  Holthoff  furnace,  where 
the  hearth  revolves,  the  Stetefeldt  shaft-furnace,  the  Edwards 
and  the  Merton  furnaces. 

In  these  various  furnaces  advantage  is  taken  of  the  heat 
developed  as  the  result  of  oxidation  of  the  sulphides  at  the  tem- 
perature of  the  furnace.  If  the  percentage  of  sulphur  is  high, 
this  is  often  enough  to  supply  the  required  heat  (after  having 
once  been  started),  without  the  aid  of  extraneous  fuel.  The 
various  mechanical  roasters  roast  ores  very  cheaply,  but  for  ores 
containing  lead,  which  may  agglomerate,  those  which  are  stirred 
with  rabbles  do  not  give  the  satisfaction  of  hand-reverberatory 
roasters.  With  a  slight  accession  of  heat  above  the  normal,  due 
to  a  variation  in  the  firing,  the  ore  is  liable  to  agglomerate  or 
stick  to  the  hearth.  In  the  hand-reverberatory  roaster,  the  same 
thing  may  occur,  but  the  furnace  in  this  case  is  quite  accessible 
to  cutter-bars  by  which  such  obstruction  can  be  removed,  and, 
even  if  the  hearth  builds  up,  the  furnace  can  still  be  used.  An 
effort  has  been  made  to  remove  these  accretions  in  the  rabble 
furnace  by  putting  in  a  flat  bar  of  iron  in  place  of  the  blade, 
stout  enough  to  tear  loose  these  accretions.  It  has  not,  however, 
proved  successful.  The  hand-reverberatory  works  to  better 
advantage  on  ores  which  need  a  high  finishing  heat  for  break- 
ing up  the  sulphates. 

The  long-hearth  reverberatory  roaster  or  calciner.  Its  essen- 
tial features  are  a  floor,  or  hearth,  heated  by  a  fire  contained  in 
a  'fire-box/  with  a  space  beneath  the  grate  called  an  'ash-pit/ 
and  separated  by  a  'fire  bridge'  from  the  hearth,  upon  which  lies 
the  ore  spread  out  over  its  entire  surface.  The  whole  is  covered 


74  THE    METALLURGY 

by  a  flat  arch  or  'roof  against  which  the  flame  'reverberates' 
(hence  the  name  of  the  furnace),  heating  the  charge  on  its  pas- 
sage from  the  fire-box  to  the  farther  or  'flue  end,'  and  thence  to 
the  chimney  or  stack. 

These  furnaces  are  distinguished  from  reverberatory  smelt- 
ing furnaces  by  having  a  much  smaller  grate  area,  and  by  having 
a  flat  hearth,  at  the  same  level  as  the  side  door-sills  of  the  furnace. 

Fig.  27  and  28  represent,  in  sectional  plan,  and  in  longitudinal 
sectional  elevations,  a  long-bedded  reverberatory  roaster. 
Its  inside  width,  for  convenience  in  stirring  and  for 
moving  along  the  charges,  may  be  fixed  at  14  ft., 
while  its  length  varies  according  to  the  ability  of 
the  ore  to  generate  heat  by  the  combustion  of  its  sulphur,  the  fire- 
place, without  this  assistance,  not  being  able  to  maintain  the 
requisite  temperature  beyond  32  ft.  from  the  fire-bridge.  The 
heat-generating  power  of  the  ore  depends  upon  the  percentage  of 
contained  sulphur,  including  the  loosely  held  equivalent  as  found 
in  iron  pyrite.  An  ore,  containing  no  more  than  10%  sulphur, 
would  therefore  be  well  roasted  in  a  short  furnace  of  16  ft. ; 
where  there  is  15%  sulphur  it  is  proper  to  add  another  hearth; 
a  20%  ore  would  work  rapidly  in  a  three-hearth  furnace ;  and 
where  the  ore  contains  25^  or  more  of  sulphur,  a  fourth  sec- 
tion should  be  added,  making  a  total  length  of  64  ft.  of  hearth, 
which  is  long  enough  for  any  ore  the  metallurgist  is  likely  to 
handle.  Hearths  of  greater  length  have  been  tried,  but  have 
not  been  found  satisfactory. 

This  kind  of  furnace  has  several  advantages  in  the  roasting  of 
ore,  namely,  in  starting  the  roasting  operation  at  a  low  tem- 
perature at  which  there  is  but  little  tendency  for  the  ore  to 
crust,  or  stick  together,  thus  ensuring  its  thorough  contact  with 
the  air;  the  saving  of  fuel  which  in  this  length  of  furnace  gives 
up  much  of  its  heat,  and  leaves  the  furnace  at  a  temperature  of 
about  260°  C ;  the  thorough  stirring  and  turning  resulting  from 
the  gradual  removal  of  the  charges  to  the  front  or  bridge-end 
of  the  furnace ;  a  uniform  firing,  and  an  economy  in  repairs  and 
construction  due  to  the  uniform  and  moderate  heat  of  the  rear 
portion  of  the  furnace,  where  red  brick  can  be  used.  Moreover, 
the  final  heating  near  the  bridge  can  be  well  performed  for 


OF   THE    COMMON    METALS. 


75 


BE 

SS 


/6  THE    METALLURGY 

breaking  up  sulphates  and  agglomerating  the  ore  (provided  it  is 
somewhat  fusible),  so  that  it  is  in  better  condition  for  a  blast- 
furnace operation.  It  must  be  remembered,  however,  that  in 
copper  smelting  these  pulverized  ores  are  rather  suited  to  the 
reverberatory  furnace. 

The  cast-iron  door-frames  of  this  furnace  are  set  6  ft.  apart 
and  opposite  one  another.  The  door  is  made  of  a  piece  of  sheet- 
iron  removable  by  means  of  a  'lifter.'  The  hearth  is  divided  by 
steps  into  divisions,  so  that  successive  charges  shall  not  mix  to 
the  detriment  of  the  roast,  due  to  the  mixing  of  the  ore  of  a  less 
roasted  charge  with  that  of  the  more  advanced  one.  The  furnace 
is  strongly  stayed  and  tied  by  buck-staves  and  tie-rods  to  resist 
the  expansion  due  to  the  heated  brick-work,  and  the  thrust  of 
the  covering  arch.  At  the  front  wall  of  the  fire-box,  openings 
of  2.]/2  by  4  in.  are  often  left  for  the  admission  of  air  above  the  fire 
level.  The  bridge  also  has  a  transverse  passage  for  cooling  it, 
and  for  conducting  air  to  the  charge  by  means  of  transverse 
ports  or  openings,  also  2.^/2  by  4  in.,  by  which  air  gets  in  through 
the  bridge-wall  to  the  hearth.  These  openings  furnish  air  which, 
together  with  that  entering  through  the  fire,  produces  an  oxidiz- 
ing atmosphere  upon  the  hearth.  The  fire-box  and  the  hearth 
for  the  first  15  ft.  should  be  of  fire-brick.  A  better  quality  of 
common  brick  may  be  used  for  the  rest  of  the  furnace,  both 
for  the  roof  and  for  paving  the  hearth.  All  such  brick,  however, 
must  be  laid  in  clay,  not  in  lime-mortar. 

The  proper  coal  for  a  roaster  is  a  free-burning,  semi-bitumi- 
nous coal  which  should  be  burned  upon  the  grate  in  a  shallow 
bed  from  6  to  8  in.  thick.  It  should  be  fed  at  intervals  of  15  to  30' 
minutes  in  small  quantities,  and  a  roaster  will  consume  5,000  to 
6,000  Ib.  in  24  hours.  With  a  roaster  of  a  capacity  of  12  tons 
daily  this  is  equal  to  a  consumption  of  at  least  20%. 

The  stack  proposed  by  Peters  for  furnishing  the  draft  to 
two  roasters,  is  42  in.  square  (internal  dimensions)  and  65  ft. 
high.  Its  cost,  together  with  that  of  the  short,  connecting  flues 
for  the  furnaces,  he  gives  at  $728,  while  the  cost  of  the  roaster 
is  given  at  $2.713. 

The  cost  of  roasting  a  copper  sulphide  in  the  long-bedded 
reverberatory  furnace  is,  according  to  Peters,  $1.81  per  ton  of 


OF    THE    COMMON    METALS. 


78  THE    METALLURGY 

ore  charged.  If  the  ore  contain  lead,  which  makes  it  more  diffi- 
cult to  roast,  the  price  may  go  up  to  $2.25  per  ton. 

The  Slagging-reverberatory  roaster.  This  consists  of  an  ordi- 
nary long-bedded  roaster  to  which  has  been  added  a  slagging- 
hearth  or  fuse-box.  Certain  ores  containing  much  zinc,  and 
which  in  roasting  produce  sulphates  that  decompose  with  diffi- 
culty, are,  in  silver-lead  practice,  roasted,  and  then  dropped 
down  into  the  fuse-box  where  the  entire  material  is  melted  into  a 
slag.  By  so  doing  the  silica  present  unites  itself  to  bases  break- 
ing up  their  combinations  as  sulphates  thus — 

Zn  SO4  +  SiO,  =  Zn  SiO3  +  SO3 
and  Pb  SO4  +  SiO2  =  Pb  SiO3  +  SO3 

The  sulphuric  anhydride  escapes  in  the  fumes  while  the  fusible 
silicates  thus  formed  are  skimmed  as  slag  into  slag-pots.  This 
fused  material  can  then  be  smelted  without  flue-dust  loss. 

Fig.  29  shows  such  a  furnace,  72  ft.  long  by  17  ft.  wide.  It 
consists  of  a  long  brick  hearth  covered  with  a  low  arch  and  ter- 
minating next  to  the  fire-box  in  a  fuse-box  or  slagging-hearth. 
A  high  heat  is  carried  at  this  point,  so  that  the  fire  bridge  must  be 
protected  by  a  water  jacket  or  coil  of  water  pipes  inserted  amidst 
the  brick-work  of  the  bridge.  Ore  is  charged  at  the  flue  end  by 
means  of  a  hopper,  also  shown,  and  gradually  worked  forward 
and  roasted  as  in  an  ordinary  long-bedded  roaster.  Thus  roasted, 
it  is  passed  by  a  paddle  to  the  fuse-box,  whose  hearth  is  at  a 
lower  level.  Here  it  is  spread  out  and  melted  down. 

The  Pearce  turret  furnace.  Three  types  of  this  furnace  have 
been  developed,  the  one-deck,  the  two-deck,  and  the  six-deck 
or  hearth.  The  greater  the  number  of  hearths,  the  more  econom- 
ical is  the  furnace  as  regards  fuel  consumption  and  output  per 
furnace.  On  the  other  hand  the  structure  becomes  much  more 
complicated.  We  will  consider  for  illustration  the  type  of  the 
two-deck  furnace  (Fig.  30  and  31).  It  consists  of  two  super- 
imposed circular  hearths  heated  by  external  fire-boxes.  The 
ore  is  fed  continuously  through  a  slit  in  the  roof  of  the  upper 
hearth,  is  stirred  and  slowly  moved  forward  through  the  circuit 
of  that  hearth,  and  falls  through  a  transverse  slit  to  the  lower 
hearth  where  it  is  stirred  and  moved  forward  as  before,  and 
finally  discharged  at  a  break  or  open-space  of  the  hearth.  The 


OF    THE    COMMON    METALS. 


79 


FIG.  30.     TWO-DECK  PEARCE  TURRET  ROASTING  FURNACE  (PLAN) 


8O  THE    METALLURGY 

stirring  is  done  by  blades,  shown  in  detail  in  Fig.  32,  which 
are  attached  to  the  radial  arms.  As  will  be  noticed  in  the 
figure,  and  at  in  in  Fig.  30,  the  rabble  blades  are  set  at  an 
angle  to  their  line  of  travel.  Nearly  touching  the  hearth, 
they  pass  through  the  ore,  stirring  it,  and  while  pushing  it  aside, 
move  it  slightly  forward.  The  blades  on  the  opposite  rabble,  in- 
clined the  opposite  way,  stir  the  ridges  thus  formed.  Thus 
about  every  minute  fresh  surfaces  are  presented  to  the  action  of 
the  fire.  An  objection  has  been  urged  against  this,  as  against 
all  the  mechanically  stirred  roasters,  that  there  is  a  creeping 
forward  of  part  of  the  less  roasted  ore  to  mix  itself  with  the 
more  roasted  portion  of  the  charge.  In  hand-roasting  this  need 
not  occur. 

Each  set  of  arms  is  carried  by  a  central  hub  revolving  with  a 
central  hollow  column.  Air  under  pressure  is  forced  from  the 
central  column  through  the  rabble  arms,  thus  cooling  them,  and. 
at  the  same  time,  furnishing  air  for  oxidation.  The  inner  wall  of 
the  furnace  has  a  continuous  slot  for  the  movement  of  the 
rabble,  the  slot  being  covered  by  a  steel  tape  or  band,  moving  with 
the  rabble  arms.  The  inner  wall,  thus  cut  in  two,  is  sustained  by 
I-beams  resting  both  on  the  external  solid  wall  and  upon  the" 
central  column.  Gears  attached  to  each  set  of  arms  are  actuated 
by  pinion  bevel  gears  and  pulley  as  shown  in  Fig.  30.  The 
fire-boxes  are  supplied  by  air  from  a  fan  under  a  few  ounces' 
pressure  ('under-grate  blast'),  so  that  the  flame  enters  the 
furnace  under  pressure,  thus  neutralizing  the  inward  suction  of 
the  outer  cool  air  due  to  the  draft,  and  the  consequent  cooling  of 
the  furnace.  Suction  increases,  however,  as  the  outlet  flue  is 
approached.  Fig.  31  shows  one  of  the  fireplaces,  having  a  step- 
grate,  used  for  burning  slack  coal.  The  ash  pit  is  closed  by  tight 
iron  doors,  so  that  the  under-grate  blast  can  be  sustained.  The 
fire-box  is  supplied  with  coal  by  means  of  a  hopper  kept  constantly 
full.  To  prevent  the  too  intense  action  of  the  heat  on  the  newly 
dropped  charge,  a  curtain  arch  (Fig.  31),  deflects  and  dis- 
tributes the  flame.  At  the  open  part  of  the  circular  hearth  each 
rabble  arm  successively  emerges,  pushing  aside  a  swinging  sheet- 
iron  door.  Continuing  its  movement  it  at  once  enters  the  heated 
hearth  through  another  swinging  door. 


82 


THE    METALLURGY 


A  furnace  with  a  7-ft.  hearth  (area  1,218  sq.  ft.)  will  treat  42 
tons  of  ore  containing  35^  sulphur  and  roasting  it  down  to 
6  to  7%  with  a  consumption  of  9.1%  fuel.  The  labor  needed 
for  a  double-deck  furnace  is  no  more  than  for  a  single-deck  one, 
while  the  fuel  cost  per  ton  of  ore  is  one-half.  The  flue-dust  is 
more  than  in  the  single-deck  furnace.  In  a  multiple-deck  Pearce 
turret  furnace  the  fuel  has  been  cut  to  1.4%  of  the  weight  of  the 
ore,  but  such  a  furnace  makes  4%  flue-dust.  The  two-deck 
furnace  requires  3  horsepower  to  run  it  and  the  cost  of  roasting 
is  98c.  per  ton.  The  cost  of  the  roaster  is  $8,000. 

riG.  9. 


OETAtU    OF  RABBLES. 

FIG.  32.     DETAILS  OF  ARMS  OF  PEARCE  TURRET  ROASTING  FURNACE. 

The  Edwards  Roasting  Furnace. — Fig.  33,  34,  and  35  represent 
respectively  a  plan  (partly  in  section),  an  elevation  (partly  in 
section),  and  a  cross-section  of  the  tilting  type  of  an  Edwards 
ore-roasting  furnace,  6  by  57  ft.  hearth  dimensions.  It  is  a  single- 
hearth  reverberatory  furnace  whose  slope  can  be  varied  by  tilting 
the  furnace  more  or  less  as  desired,  but  having  in  this  case  a 
slope  of  2  in.  per  foot  toward  the  discharge  end,  which  is  near 
the  fire-box.  The  stirring  and  moving  along  of  the  charge  is 
effected  by  rabbles  fixed  to  vertical  shafts  as  shown  in  Fig.  35. 


OF   THE    COMMON    METALS.  83 

The  ore  is  fed  from  a  storage  hopper  at  a  by  a  screw  conveyor 
to  the  hearths  of  the  furnace,  is  moved  along  by  the  revolving 
rabbles  whose  circles  intersect,  and  is  discharged  through  an  open- 
ing near  the  side  of  the  hearth  at  the  fire  end  to  a  vibrating  con- 
veyor and  then  to  the  cooling-floor.  Slides  at  the  bottom  of  the 
conveyor-trough  may  be  opened  as  desired  to  regulate  the  point 
of  discharge  upon  the  cooling-floor.  The  furnace  takes  one 
horse-power  to  operate  it,  and  has  a  daily  capacity  of  25  tons, 
roasting  sulphide  of  30  to  35^  sulphur  to  from  3  to  8c/< •.  The 


PLAN  OF  EDWARDS  ROASTING  FURNACE 


IB      a   i  Q    I  B\|  jgff  EJ  I    0      a  I 
.liti-iMfliiiriiii giiiiiiiilivijigia    i     mitt I 


FIG.  34.     ELEVATION  OF  EDWARDS  ROASTING  FURNACE 


blades  or  plows  of  the  rabbles  can  be  easily  replaced  through  the 
doors  adjacent  to  the  rabbles.  The  moving  parts  are  quite  de- 
sirable, and  the  furnace  has  proved  efficient  in  practice.  Large 
furnaces  of  the  duplex  type,  120  by  12  ft.,  have  been  built,  having 
two  rows  of  shafts  and  rabbles  with  a  capacity  of  60  tons  daily ; 
these  furnaces  have  a  fixed  hearth. 

The  Wethey  Roasting  Furnace. — Fig.  36  is  a  perspective  view 
of  the  furnace,  and  Fig.  37  a  cross-section  of  its  two  straight 
hearths.  The  roasting  is  done  on  the  upper  reverberatory  hearth, 
and  the  lower  one  is  for  cooling  the  ore  after  it  has  been  roasted. 


84 


THE    METALLURGY 


The  upper  hearth,  121  by  12  ft.,  is  held  firmly  between  two  heavy 
I-beams  suspended  from  the  main  frame  so  as  to  leave  a  slit  in 
each  wall  along  which  the  carriages  travel.  These  carriages  or 
rabbles  are,  at  each  side,  attached  to  an  endless  chain  so  that 
they  can  be  dragged  through  the  hearth,  stirring  and  gradually 
advancing  the  ore.  They  return  by  the  lower  hearth,  upon  which 
the  ore  from  the  upper  one  slides  by  a  chute,  and  wrhere  it  is 
moved  along,  cooled,  and  discharged  to  a  screw-conveyor  that 
delivers  it  to  the  mill  for  further  treatment.  To  hasten  the  cool- 
ing, so  that  it  may  not,  when  cool,  injure  the  conveyor  and  the 


FIG.  35.     CROSS-SECTION  OF  EDWARDS  ROASTING  FURNACE. 

elevator  to  the  mill,  water-cooled  pipes  are  laid  the  length  of  the 
hearth  in  grooves  between  its  brick  paving  and  flush  with  their  top 
surface.  The  ore  is  constantly  fed  into  the  furnace  at  the  driving 
end  (the  right  in  the  perspective  view)  and  is  heated  by  three 
fire-boxes  alternately  set  on  either  side  of  the  furnace.  The  flame, 
entering  from  them  by  the  flue  through  the  roof  and  descending 
toward  the  hearth,  moves  horizontally  to  the  exit  in  the  stack  in 
the  roof  near  the  feed  end.  Thus  the  ore  and  the  flame  move  in 
opposite  directions.  The  endless  chains  pass  around  the  sheaves 
at  the  ends  of  the  furnace,  and  around  sprocket-sheaves  at  the 
driving  end  that  impel  them.  The  stirring  plows  or  blades  are 
set  at  opposite  angles  upon  the  rabbles,  so  that  the  end  thrusts 
are  balanced.  The  rabbles  are  so  arranged  that  they  can  easily 


OF    THE    CO'MMON    METALS. 


FIG.  36.     PERSPECTIVE  VIEW  OF  WETHEY  ROASTING  FURNACE. 


'r; 


1 


FIG.  37.     CROSS-SECTION  OF  WETHEY  ROASTING  FURNACE. 


86  THE    METALLURGY 

be  removed  without  disturbing  the  carriage  connections.  There 
are  four  carriages  and  they  pass  through  the  hearth  at  the  rate 
of  loo  ft.  per  min.  There  are  sheet-iron  flap-doors  hinged  on  the 
top  edge  at  each  end  of  the  roasting-hearth  (as  in  the  Pearce 
turret- furnace).  These  are  ordinarily  closed,  but  are  lifted  by 
the  stirrer  carriage  in  passing  into  and  out  of  the  upper  hearth. 
It  might  be  thought  that  the  slits  at  the  sides  of  the  furnace  would 
admit  too  much  air,  and  also  injure  the  draft,  but  the  furnace 
is  found  to  work  well  notwithstanding. 

The  Evans-Klcpctko  furnace.  This  belongs  to  the  type  of 
McDougall  furnaces,  and  is  shown  in  Fig.  39.  It  is  a  ver- 
tical cylindrical  furnace  with  arched  horizontal  hearths,  having 
drop  or  discharge  openings  for  the  ore  alternately  at  the  centre 
and  at  the  periphery  of  the  hearths.  To  stir  the  ore  and  to  move 
it  alternately  to  these  drop  openings,  there  is  a  central  revolving 
shaft,  having  horizontal  radial  stirring-arms  provided  with 
stirring-blades  set  at  an  angle  to  their  direction  of  motion. 

The  blades  of  the  arms  on  the  even-numbered  hearths  are 
inclined  opposite  to  those  on  the  odd-numbered,  so  that  on  the 
odd-numbered  they  push  the  ore  to  the  central  drop-opening,  and 
the  even-numbered  toward  the  periphery.  The  ore.  fed  con- 
tinuously to  the  furnace,  drops  upon  the  first  hearth  near  its  outer 
edge ;  the  blades  of  the  rabble-arms  of  that  hearth  stir  and  move 
it  gradually  toward  the  central  drop-opening,  where  it  passes  to 
the  second  hearth.  The  rabbles  of  this  hearth  again  stir  and 
move  it  to  the  exterior  drop-openings  where  it  falls  to  the  third 
hearth  and  so  on.  At  the  lower  hearth,  openings  are  left  by 
which  it  is  finally  dropped  into  a  hopper.  The  ore  spread  evenly 
on  the  hearth  thus  travels  zig-zag  through  the  furnace,  gradually 
passing  from  periphery  to  centre  and  rice-  rcrsa.  The  ore,  con- 
tained in  a  hopper,  is  mechanically  fed  in  a  continuous  stream  to 
the  periphery  of  the  upper  hearth.  Here,  as  it  is  moved  toward 
the  central  discharge  opening,  it  dries  out,  roasting  beginning  on 
the  second  hearth.  On  the  third  hearth  the  ore  roasts  freely  with 
the  emission  of  numerous  sparks,  and  some  sulphates  are  formed ; 
on  the  fourth  there  are  no  sparks,  and  the  ore  has  attained  its 
highest  temperature.  On  the  fifth,  the  ore  looks  less  bright,  es- 
pecially at  the  discharge,  where  it  has  begun  to  cool  off. 


OF   THE    COMMON    METALS.  »7 

The  air  for  oxidation  is  admitted  by  doors  in  the  side  of  the 
furnace,  mostly  at  the  bottom  ones.  The  gases,  and  with  them, 
the  dust,  passing  up  through  the  drop-openings,  are  drawn  off 
through  the  flues  at  the  top.  In  starting,  the  furnace  has  to  be 
heated  up  to  the  kindling-temperature  of  the  ore,  which,  if  rich 


FIG.  39.     EVANS-KLEPETKO  FURNACE   (ELEVATION). 

enough  in  sulphur,  will  burn  of  its  own  accord  without  the  addi- 
tion of  any  fuel.  If  low  in  sulphur,  additional  heat  must  be 
supplied  by  one  or  more  external  fireplaces  located  near  the 
bottom  of  the  furnace.  To  protect  the  rabble-arms  from  the 
intense  action  of  the  fire,  they  are  water-cooled.  The  cooling 


88  THE    METALLURGY 

water  is  forced  down  to  near  the  bottom  of  the  9-in.  hollow  cen- 
tral shaft  in  a  3~in.  pipe  and  out  to  the  ends  of  the  horizontal 
rabbles  or  stirring  arms  in  i-in.  pipes.  It  then  returns  up 
the  annular  space  between  the  3-in.  pipe  and  the  hollow  shaft, 
and  discharges  at  the  top  through  two  spouts  into  a  launder.  The 
furnace  is  18  ft.  3  in.  high  and  15  ft.  10  in.  diam.,  having  six 
hearths.  The  shell  is  of  fain,  boiler  iron,  and  is  lined  with  9 
inches  of  brick.  Each  hearth  has  two  stirring-arms  making  one 
revolution  per  minute. 

A  furnace  treats  in  24  hours  40  tons  of  sulphide  ore  of  35% 
sulphur,  reducing  it  to  7%.  About  4%  of  flue-dust  is  made,  and 
the  roasted  ore  itself  contains  more  ferric  oxide  and  is  lighter 
and  more  porous  than  would  be  the  case  with  the  product  of  the 
hand  reverberatory-roaster.  The  cost  of  roasting  for  the  above- 
named  ore  may  be  given  at  35c.  per  ton.  the  lowest  thus  far 
known  for  any  furnace.  The  compact  form  of  the  furnace,  re- 
ducing radiation  to  a  minimum,  enables  it  to  roast  with  the  con- 
sumption of  little  or  no  fuel.  Taking  its  capacity  into 
consideration,  it  is  of  moderate  price  and  is  easy  to  repair. 

The  White-Hoi(.'ell  Roasting  Furnace.  Fig.  40  is  an  elevation 
of  this  furnace,  which  has  a  continuous  feed  and  consists  of  a 
cylinder  5  ft.  external  diameter  by  32  ft.  long,  slightly  inclined, 
supported  on  friction  rollers  and  revolved  between  a  stationary 
fire-box  at  one  end,  and  a  dust-chamber  and  flue  connected  to 
the  stack  at  the  other.  That  portion  of  the  cylinder  near  the 
fire  is  of  larger  diameter  to  permit  of  it  being  lined  with  fire- 
brick where  needed,  leaving  the  interior  of  the  same  diameter 
throughout.  For  projecting  fire-brick  ridges  serve  to  raise 
the  ore  as  the  cylinder  revolves  and  to  shower  it  back  through 
the  flame  so  as  to  more  rapidly  roast  it.  The  unlined  part  for 
the  same  reason  is  provided  with  longitudinal  cast-iron  shelves. 
The  furnace  is  fed  at  the  flue  end  by  means  of  a  screw-feed,  and 
as  the  cylinder  revolves,  the  ore  works  along  to  its  lower  end, 
passing  out  between  the  cylinder  and  the  fire-box  to  a  vault  or 
brick  chamber  whence  it  can  be  withdrawn  when  cool.  The 
furnace  has  been  chiefly  used  for  chloridizing  roasting. 

The  Bruckner  roasting  furnace.  This  furnace  consists  of  a 
horizontal  cylinder  of  plate-steel  lined  with  brick,  and  revolving 


OF   THE    COMMON    METALS. 


89 


9O  THE    METALLURGY 

between  a  fire-box  and  a  flue.  The  flame  from  the  fire-box  is 
drawn  directly  through  the  cylinder  to  the  flue.  The  cylinder  is 
provided  with  man-holes  for  charging  and  discharging  the  ore. 
and  is  served  by  a  double  hopper,  large  enough  to  hold  a  full 
charge. 

Fig.  41  is  a  perspective  view  of  the  furnace,  8  ft.  6  in.  diam.  by 
28  ft.  long. 

The  cylinder  is  truncated  at  the  ends  for  the  easier  discharge 
of  the  ore,  and  constricted  at  the  inlet  and  outlet  openings 
intended  for  the  passage  of  the  flame  and  gases.  The  present 
practice  is  to  revolve  the  cylinder  slowly,  say  once  an  hour,  since 
its  contents,  even  at  this  speed,  are  constantly  shifting:  the  older 
way  was  to  revolve  it  once  in  four  minutes.  Motion  is  commu- 
nicated through  spur-gearing  to  a  circular  spur-gear  rack  on  the 
cylinder,  and  the  slow  motion  is  attained  preferably  through 
worm  and  worm-gear.  At  the  time  of  discharging,  however,  the 
cylinder  must  be  revolved  quickly  in  order  to  remove  the  ore 
promptly.  Charges  which  contain  lead  are  liable,  with  a  slight 
accession  of  heat  over  the  normal,  to  agglomerate  slightly  and 
'hang  up' :  that  is,  the  ore  attaches  itself  in  a  layer  to  the  brick 
lining.  Should  this  occur,  the  movable  fire-box  may  be  pushed 
aside  and  the  layer  removed  with  long  slicer-bars.  The  attach- 
ment is  so  slight  that  when  the  layer  is  cut  through  even  in  one 
line  from  end  to  end  of  the  cylinder,  thus  destroying  its  key,  the 
rest  falls  off  as  the  cylinder  again  revolves,  and  the  whole,  again 
breaking  up,  is  ready  for  further  roasting.  A  slight  cohesion  of 
the  finer  particles  need  not  interfere  with  obtaining  a  good  roast. 

The  charge  having  been  dropped  into  the  furnace  and  the 
man-hole  openings  closed,  it  is  vigorously  fired  upon  in  order  to 
get  the  ore  burning  by  its  own  oxidation. 

This  takes  about  six  hours,  the  ore  slowly  becoming  of  a 
visible  red  heat,  first  at  the  fire-box  end  and  finally  at  the  flue 
end.  The  charge  thus  started  burns  by  its  own  heat  for  12 
hours  more,  the  fire  having  meanwhile  been  withdrawn  from  the 
fire-box.  Indeed,  the  action  is  so  vigorous  that  the  charge  must 
be  watched  and  the  admission  of  air  limited,  for  fear,  in  case  of 
leady  ores,  of  agglomeration.  As  the  heat,  at  the  expiration  of 
this  stage,  slacks  off,  firing  with  coal  is  resumed,  gradually  in- 


OF    THE    COMMON    METALS. 


92  THE    METALLURGY 

creasing  the  heat  to  the  finish.  In  this  way  sulphates  are  de- 
composed, and  the  oxidation  of  the  charge  completed.  The 
whole  operation,  in  the  case  of  lead-bearing  ores,  takes  48  hours — 
sometimes  longer.  In  the  case  of  copper-bearing  ores  which  do 
not  agglomerate,  and  which  need  not  be  roasted  so  closely,  24  to 
36  hours  is  sufficient.  In  the  former  case,  the  sulphur  remaining 
may  be  from  3  to  5%,  in  the  latter  7  to  8%.  The  ore  is  dis- 
charged by  opening  all  the  man-holes  and  revolving  the  cylinder 
quickly.  At  each  half-revolution,  part  of  the  charge  runs  out, 
until,  in  5  to  10  minutes,  most  of  it  is  removed.  The  little  left 
behind  mixes  with  the  next  charge.  The  hot  escaping  ore  is  re- 
ceived into  large  charge-cars,  transferred  and  dumped  into  bins  to 
cool.  All  the  man-holes  save  two  are  closed,  these  latter  being 
brought  under  the  hopper-spouts.  The  hopper-slides  are  pulled 
and  the  ore  speedily  runs  into  the  cylinder.  On  closing  the  man- 
holes, the  charge  is  ready  for  firing  upon.  These  operations  of 
discharging  and  charging  need  take  no  more  than  20  minutes. 
A  large  cylinder  8  ft.  6  in.  by  28  ft.  will  take  a  charge  of  30  tons, 
which  makes  a  capacity  of  15  tons  daily  for  leady  sulphides,  and 
of  20  to  30  tons  for  copper-bearing  sulphides.  The  cost  of  roast- 
ing, for  leady  ores,  is  85c.  per  ton ;  for  copper  sulphides  roasted 
to  6  or  7%,  but  42c.  per  ton.  The  cost  of  a  Bruckner  cylinder  may 
be  given  at  $3,000. 

The  Stctcfeldt  Roasting  Furnace. 

This  furnace  (Fig.  42)  consists  of  a  vertical  brick  shaft,  B, 
some  25  ft.  high,  having  fire-boxes,  G  G,  near  the  bottom  and 
a  flue  opening  near  the  top.  By  means  of  a  screen  feeder,  A, 
pulverized  ore  is  continuously  sifted  into  the  shaft  and,  during 
the  few  seconds  it  is  falling  through  the  heated  air,  becomes 
roasted.  The  flue-dust  passing  off  through  H  is  mostly  caught 
in  the  hoppers,  FF,  and  the  remainder  in  the  main  flue  forming 
a  continuation  of  the  flue,  D,  and  terminating  in  a  chimney  or 
stack.  An  auxiliary  fire-box,  E,  completes  the  roasting  of  the 
flue-dust,  which  may  amount  to  30^  of  the  ore  fed  in.  The 
furnace  has  a  capacity  of  40  tons  in  24  hours. 


OF    THE    COMMON    METALS. 


93 


FIG.  42.    STETEFELDT  ROASTING  FURNACE. 


94  THE    METALLURGY 

13.       POT-ROASTING   OF   ORES. 

We  have  already  noted  the  difficulty  of  roasting  galena 
ores,  but,  by  one  of  the  pot-roasting  methods,  galena  may 
be  freed  from  most  of  its  sulphur.  In  these  processes 
the  moistened  sulphide  ore,  carrying  galena  mixed  with  a  certain 
proportion  of  limestone  and  silicious  ore.  is  charged  into  a  large 
hemispherical  cast-iron  pot,  usually  of  a  capacity  of  from  8  to 
10  tons.  Air,  under  low  pressure  (one  to  four  ounces)  is  in- 
troduced under  the  false  bottom  of  the  pot  and  is  blown  through 
the  charge ;  a  little  wood  fire,  if  necessary,  or  hot  ore,  being  at 
first  put  under  it.  An  exothermic  reaction  takes  place,  the  charge 
becoming  red  hot  and  S(X  and  SO3  coming1  off.  At  the  end  of  the 
reaction,  sometimes  extending  to  16  hours,  desulphurization  is 
almost  completely  effected.  The  pot  is  then  inverted  and  the  solid, 
red-hot  mass  dumped  out.  This  is  then  broken  up  to  a  size  suitable 
for  the  blast-furnace.  Owing  to  its  lump  form,  it  permits  the  air 
of  the  blast-furnace  to  come  through  more  readily,  thus  increas- 
ing tonnage,  while,  owing  to  the  low  percentage  of  sulphur,  the 
production  of  matte  is  greatly  decreased.  Besides  this,  since  the 
temperature  of  desulphurization  is  not  high,  the  loss  of  silver 
and  of  lead  is  reduced  as  compared  with  that  in  an  agglomera- 
ting-roasting. 

There  are  three  patented  variations  of  the  process.  In  the 
first,  the  Huntington-Heberlein,  the  ore  mixed  with  limestone  is 
partially  roasted  down  to  12  to  i4f/r  S,  and  charged  to  the 
pot  while  hot,  so  that  no  extraneous  fuel  is  needed  to  start  the 
reaction.  In  the  second,  the  Savelsberg,  the  ore,  mixed  with  the 
proper  proportion  of  limestone  and  of  silicious  ore,  is  charged 
directly  to  the  converter,  there  being  no  preliminary  roasting.  In 
the  third,  the  Carmichael-Bradford  process,  the  ore  is  mixed  with 
a  proportion  of  gypsum,  and  then  charged  directly  to  the  pot. 

14.     COST  OF  ROASTING. 

Ore-roasting  in  heaps  at  Jerome,  Arizona,  costs  8oc.  per  ton, 
including  general  expenses. 

Ore-roasting  in  stalls  costs  54C.  per  ton.  A  battery  of  roasting 
stalls,  each  8  by  8  ft.  in  size  would  cost  $3,300,  to  which  should  be 


OF    THE    COMMON    METALS.  95 

added  $800  for  the  cost  of  the  stack  (common  to  all  the  stalls), 
making-  a  total  of  $4,100.  Such  a  plant  has  a  life  of  about  six 
years. 

In  rci'crbcrator\  roasting  in  long-bedded  hand-rabbled  furnaces 
the  lowest  price  when  roasting  copper  ores  is  $1.25  per  ton.  For 
roasting  lead-bearing  ores  $1.75  per  ton  is  a  moderate  figure. 
The  cost  of  erecting  such  a  roaster  may  be  given  at  $3,130,  in- 
cluding that  of  one-half  of  the  stack  needed  for  two  roasters  in 
common. 

The  Allen-O'Hara  automatic  furnace,  having  two  hearths  each 
94  by  9  ft.  will  roast  45  to  50  tons  daily  to  8'/  S  at  a  cost  of  /8c. 
per  ton. 

The  Wethey  furnace,  having  four  hearths  each  65  by  10  ft., 
will  roast  90  tons  daily  to  5  or  6r/(  S  at  a  cost  of  6oc.  per  ton. 

The  Pearce  double-deck  furnace  (7- ft.  hearth)  \vill  roast  42 
tons  daily  to  6c/<  S  at  a  cost  of  o8c.  per  ton. 

The  Herreshoff  furnace,  having  five  hearths  each  10  ft.  10  in. 
diam.,  will  roast  40  tons  daily  to  7c/(  S  at  a  cost  of  5oc.  per  ton 

The  McDougall-Evans-Klepetko  furnace,  with  six  hearths 
each  15  ft.  10  in.  diam.,  will  roast  40  tons  daily  to  7*/r  S  at  a 
cost  of  35c.  per  ton. 

The  Bruckner  cylinder  roaster,  8^2  ft.  diam.  by  22  ft.  long,  will 
roast  TO  tons  daily  to  4^  S  at  a  cost  of  8oc.  per  ton. 

It  is  to  be  noticed  that  the  low  cost  of  roasting  on  some  of 
these  furnaces  is  due  to  their  using  no  fuel  after  they  are  fully  in 
operation  Such  furnaces  have  several  hearths,  are  compact,  and, 
on  account  of  that  compactness,  lose  but  little  heat  by  radiation. 


PART  III.    GOLD 


PART  III.     GOLD. 

15.     GOLD  ORES. 

Gold  occurs  in  nature,  both  native  and  combined  with  tel- 
lurium. 

Native  gold  occurs  in  rock  disseminated  through  it  in  grains  or 
particles  of  varying  size.  It  is  found  not  only  in  quartz,  but  in 
hematite,  iron  pyrite,  arsenical  pyrite,  blende,  and  galena.  In 
pyrite  it  may  occur  as  films  on  the  faces  of  the  crystals.  Such 
rock  in  place  is  called  gold  rock  or  reef  rock. 

When  these  rocks  have  been  broken  up  by  alluvial  action  the 
particles  of  gold  are  swept  along  by  the  water  mixed  with  sand 
and  gravel,  and  are  deposited  in  beds,  the  pebbles  and  boulders 
being,  in  general,  barren  of  gold.  Gold  occurring  in  this  way 
is  called  alluvial  gold,  and  is  recovered  by  methods  of  hydraulic 
mining  or  dredging,  which  belong  rather  to  mining  engineering 
than  to  metallurgy.  We  will  consider,  therefore,  the  treatment 
of  gold  rock  or  ore  in  place. 

Gold  Tellurides. — In  South  Dakota,  at  Cripple  Creek,  Colo., 
and  in  West  Australia  is  to  be  found  gold  combined  with  tel- 
lurium in  the  form  of  sylvanite  and  petzite  (Au,  Ag)  Te2  and 
as  calaverite  (Au  Te2)  the  latter  having  41.4%  Au  to  $7.3%  Te. 

l6.      GOLD-MlLLING  AND  AMALGAMATION. 

Free-milling  gold  ores,  in  which  the  ore  is  oxidized,  can  be 
well  and  cheaply  treated  by  milling  the  ore  and  amalgamating  the 
crushed  pulp,  while  the  tailing  or  residue,  left  after  such  extrac- 
tion, is  run  to  waste.  If,  however,  the  ore  contains  some  pyrite, 
not  all  the  gold  content  can  be  recovered  so  simply.  The  pyrite, 
though  being  heavy,  is  caught  on  concentrating  tables  and  the  con- 
centrate is  then  sent  away  to  be  smelted  or  to  be  further  treated  by 
leaching  methods.  At  Treadwell  Island,  Alaska,  about  half  the 


IOO  THE    METALLURG\ 

gold  is  obtained  by  amalgamation,  and  the  concentrate,  one- 
fortieth  of  the  total  weight  of  the  ore,  contains  the  remainder  of 
the  gold. 

Stamp-mill  Amalgamation. — This  consists  in  crushing  the  gold- 
bearing  rock  in  a  stamp-mill  and  in  some  instances,  feeding 
mercury  in  small  quantities  (about  1.5  oz.  per  ounce  of  con- 
tained gold)  into  the  mortar  while  the  crushing  proceeds.  By 
regulating  the  water  supply  and  the  height  of  discharge  from 
the  mortar,  the  mercury  is  kept  distributed  throughout  the  ore 
pulp,  as  the  mixture  of  water  and  finely  ground  ore  is  called.  By 
the  fall  of  the  stamps  the  gold  is  brought  into  intimate  contact 
with  the  mercury,  and  forms,  with  the  gold,  an  amalgam  which 
settles  at  the  bottom  of  the  mortar  between  the  dies,  and  also 
becomes  attached  to  amalgamated  copper  plates  (inside  plates) 
lining  the  sides  of  the  mortar.  As  all  the  gold,  how- 
ever, is  not  amalgamated  within  the  mortar,  the  issuing  pulp 
flowing  over  a  flat  amalgamated  table  or  apron  plate  (4  ft.  wide 
by  5  to  6.5  ft.  long),  and  is  further  acted  upon,  the  gold  attaching 
itself  to  the  plate,  together  with  any  mercury  or  amalgam  that 
may  escape  from  the  mortar. 

The  Battery. — Fig.  43  is  a  perspective  view  of  a  lo-stamp  bat- 
tery used  in  gold  milling. 

In  Fig.  43,  the  parts  are  as  follows :  A,  mortar  block ;.  B,  mud 
sills ;  C,  cross-sills ;  D,  posts ;  E,  platform ;  F,  G,  buckstaves ;  H, 
lower  guide  timbers ;  /,  upper  guide  timbers ;  /,  mortar ;  K, 
screen ;  L,  die ;  M,  shoe ;  N,  boss  or  head ;  O,  stem ;  P,  tappet ;  R, 
cam  shaft ;  S,  collars ;  T,  cam  shaft  boxes ;  U,  cams ;  V,  cam 
shaft  pulley ;  W,  line  shaft ;  X,  tightening  pulley ;  Y,  water  pipes ; 
Z,  automatic  feeder. 

On  the  cross-sills  is  shown  the  frame  and  amalgamated  apron- 
plate,  which  in  the  cut  is  omitted  from  the  five  nearer  stamps. 

The  mortar  block.  A  is  often  made  of  timbers  set  on  end  as 
shown  in  Fig.  44  and  45.  These  timbers  are  set  in  a  pit  going 
down  to  the  solid  rock  or  on  a  concrete  foundation,  as  upon 
their  solidity  depends  the  durability  of  the  mill.  They  are  bolted 
together  horizontally,  and  are  held  vertically  by  a  tamping  of 
sand,  rock,  or  concrete,  which  fills  the  pit.  This  block  or  foun- 
dation is  also  made  of  concrete  which  is  cheaper  and  more 


OF    THE    COMMON    METALS. 


IOI 


durable.  On  the  mortar-block  is  firmly  bolted  with  long  holding- 
down  bolts,  the  mortar  /  with  three  thicknesses  of  blanket,  or  a 
piece  of  heavy  belting  or  rubber  sheeting,  interposed  to  give  an 
even  bearing. 


FIG.  43.     PERSPECTIVE  VIEW  OF   10-SxAMP   BATTERY. 


The  stamp-frames,  consisting  of  the  mud  sills  B,  the  cross-sills, 
C,  posts  D,  braces,  buckstaves  G,  H,  and  guide-timbers  H  and  /, 
are  of  dimensions  given  in  Fig.  44  and  45.  The  mud-sills  run 
the  length  of  the  mill  and  carry  also  the  line-shaft.  To  the  guide- 


102 


THE    METALLURGY 


OF   THE    COMMON    METALS. 


103 


timbers  are  bolted  the  guides,  commonly  made  of  two  planks,  each 
4  by  14  in.,  bored  with  holes  3  in.  diam.  for  the  wrought  iron 
or  steel  stems. 


12  I- 4-9" H£  (• 4-9' »| \Z 


* 

u 

,...^ 

o«' 

•       •4- 

9~         » 

£9-' 

I 

) 

£ 

) 

FIG.  45. .     FRONT  ELEVATION  OF  K)-STAMP  BATTERY. 

There  are  two  kinds  or  methods  of  framing,  called  A-frames 
and  knee-frames.  Both  Fig.  43  and  44  are  of  the  knee-frame 
type.  The  A-frame  is  braced  by  an  inclined  brace  to  the  cross- 
sills  and  tied  by  long  tie-rods  to  the  same.  The  A-frame  is 
suited  to  lighter  stamps ;  the  knee-frame  to  heavier  stamps,  as 


104 


THE    METALLURGY 


being  heavier  and  more  solid.     Stamp-frames  are  also  made  of 
steel. 

The  mortars  are  boxes  of  cast  iron  containing,  at  one  side, 
feed  openings  where  the  ore  enters,  screen  openings  on  one  or 
both  sides  for  screens  through  which  the  stamped  ore  or  pulp 


I  ' 

FIG.  46.     SINGLE-DISCHARGE   MORTAR. 

discharges,  a  base  to  receive  the  dies  and  sides  to  which  are  often 
fastened  inside  amalgamated  plates.  Two  designs  of  mortars  are 
shown  that  weigh  2  to  3  tons  each.  Fig.  46  illustrating  the 
single-discharge  type  for  gold-milling,  and  Fig.  47  a  double- 
discharge  mortar  for  silver-milling.  Upon  both  mortars  will  be 


OF    THE    COMMON    METALS. 


105 


noticed  a  deflecting  lip  terminating  the  feed  opening,  and  so 
arranged  as  to  discharge  the  ore  well  to  the  centre  of  the  mortar, 
and  (in  the  case  of  Fig.  46)  protecting  the  inside  plate,  if  any, 
from  the  direct  wear  of  the  ore.  Screen-openings  are  provided, 
both  at  the  front  and  back  of  the  mortar.  The  screen  is  mounted 


FIG.  47.     DOUBLE-DISCHARGE   MORTAR. 


in  a  wooden  frame  and  is  made  of  wire-cloth  or  of  punched-plate, 
the  holes  in  the  latter  case  being  either  slotted  or  round. 

Of  the  two,  the  punched  steel  plate  has  the  advantage  for 
strength  and  cheapness  of  first  cost.  On  the  other  hand,  brass 
wire-cloth  gives  increased  effective  discharge  area ;  thus  in  the 


IO6  THE    METALLURGY 

case  of  the  Xo.  7  punched  plate,  above  given,  we  have  an  effective 
discharge-opening  of  but  10%  of  the  total  area ;  while  in  the  case 
of  the  wire  screen,  it  is  27%  and,  in  consequence,  less  sliming  of 
the  ore  occurs  because  of  its  prompter  escape  from  the  battery. 

The  chuck  block. — To  increase  the  height  of  discharge,  which 
is  the  vertical  distance  from  the  top  of  the  die  to  the  bottom  of 
the  screen  opening,  a  block  is  put  into  the  discharge-opening 
below  the  screen,  as  shown  in  Fig.  46  and  47.  As  shown  in  Fig. 
46,  an  inside  plate  is  attached  to  this  as  well  as  to  other  parts  of 
the  interior  of  the  mortar  for  the  purpose  of  collecting  as  much 
of  the  gold  as  possible  within  the  mortar. 

Fig.  46  and  47  show  the  foot-plates  which  cover  the  bottom  of 
the  mortar,  and  upon  which  rest  the  five  cylindrical  dies  8  to  9 
in.  diam.  by  3^  to  7  in.  high  when  new.  The  connection  of  the 
stamp-shoe,  head  or  boss,  and  stamp-stem  are  also  clearly  shown. 
In  inserting  a  shoe,  wood  shims  are  placed  around  its  shank,  and 
the  boss  with  stem  affixed,  dropped  upon  it,  thus  holding  it  in 
place  by  friction. 

Dies  and  shoes  are  made  of  chilled  cast  iron  and  of  manganese 
or  chrome  steel,  chilled  cast  iron  being  quite  common.  The 
wear  of  shoes  and  dies  is  from  0.25  to  1.25  Ib.  per  ton  of  ore 
crushed,  according  to  the  toughness  of  the  ore.  In  the  case  of 
chilled  cast-iron  0.75  Ib.  at  6c.  per  Ib.,  or  4.5c.  per  ton,  may  be 
given  as  an  average  figure. 

The  reversible  tappets,  P ,  Fig.  44,  of  cast-iron,  are  keyed  to  the 
stems  and  acted  on  by  the  cams  U  which  lift  them  from  8  to  16 
in.,  according  to  the  drop  required,  at  the  same  time  giving  the 
stem  a  partial  rotation  on  its  axis  to  obtain  an  even  wear  of 
the  shoe  and  die.  The  replacing  of  a  cam  of  the  ordinary  type 
is  a  tedious  operation,  and  to  overcome  this,  a  self-tightening  cam 
such  as  the  Canda  has  been  devised,  as  shown  in  Fig.  49.  It 
consists  of  a  curved  tapering  key,  a,  which  is  fitted  to  a  tapering 
recess  in  the  cam  b.  Upon  the  shaft  c  is  set  a  pin,  which  engages 
in  a  recess  in  the  key,  so  that,  when  cam  and  key  are  slipped  on 
the  shaft  over  the  pin.  and  then  turned,  the  key  wedges  itself, 
securely,  uniting  the  cam  to  the  shaft.  It  will  be  noticed  that  the 
cams  are  set  upon  the  shaft  at  various  angles  so  that  the  stamps 
drop  at  regular  intervals  and  in  a  predetermined  order.  The 


OF    THE    COMMON    METALS. 


107 


stamps,  being  numbered  consecutively,  counting  from  left  to 
right  in  a  five-stamp  battery,  a  favorite  order  of  drop  is  No.-i,  4, 
2,  5  and  3.  When  it  is  desired  to  stop  any  one  of  the  stamps 
it  may  be  'hung  up'  by  putting  under  the  tappet  the  finger  L, 
Fig.  44. 

17.    OPERATION  OF  A  STAMP  BATTERY. 

From  the  storage  bin  (see  Fig.  44)  the  ore  passes  by  a  chute  to 
the  automatic  feeder.  At  each  down-stroke  of  one  of  the  stamps 
a  collar  on  it  strikes  the  end  of  a  horizontal  lever  there  shown, 
slightly  revolving  the  feed-plate  of  the  feeder,  so  that  some  ore 
drops  into  the  feed-opening  of  the  mortar.  As  ore  gets  under  the 


FIG.  48.       PUNCHED   SCREEN. 


FIG.  49.    CANDA  CAM. 


shoe,  the  stroke  of  the  stamp  shortens,  the  horizontal  lever  is  less 
moved,  and  the  feed  is  correspondingly  lessened,  thus  acting 
automatically  according  to  the  needs  of  the  battery,  In  wet- 
stamping,  water  is  also  fed  in,  which,  mixing  with  pulverized  ore, 
forms  a  pulp  which  is  splashed  out  through  the  mortar-screen. 

California  and  Colorado  practice  in  gold-milling.  For  free- 
milling  gold  ores,  where  the  gold  is  coarsely  knit  in  the  ore,  and 
where,  consequently,  fine  grinding  is  not  needed,  we -prefer  the 
California  practice,  by  which  large  capacity  is  attained.  On  the 
other  hand,  in  Gilpin  County,  Colo.,  the  ore  contains  10%  of  a 
gold-bearing  pyrite,  in  which  the  gold  is  finely  disseminated.  In 
order  to  grind  this  fine,  it  must  be  longer  retained  within  the 


IO8  THE    METALLURGY 

* 

mortar.  \Yhile  thus  retained,  the  fine  float-gold  set  free,  has  time 
to  come  in  contact  with  the  inside  amalgamated  plates  by  which 
some  75%  is  caught.  In  California  practice,  the  stamps  weigh- 
ing 900  to  1,100  Ib.  each,  drop  80  to  no  times  per  minute,  with 
a  fall  of  from  5  to  9  in. ;  in  Colorado  practice  the  stamps,  weigh- 
ing 600  to  800  Ib.  each,  run  at  26  to  30  drops  per  minute,  with 
a  1 6  to  2O-in.  drop.  As  a  general  principle,  the  greater  number 
of  drops  per  minute,  the  greater  the  tonnage  put  through,  but 
this  is  also  increased  by  a  high  drop.  To  grind  the  ore  more 
finely,  and  to  retain  it  within  the  battery  for  amalgamation,  a 
high  discharge  is  needed.  This  is  attained  by  putting  in  a  higher 
chuck-block.  This  height  in  Colorado  practice  amounts  to  from 
13  to  1 8  in.  from  the  top  of  the  die  to  the  lower  edge  of  the 
screen  opening.  In  California  practice  the  height  of  discharge  is 
but  5  in.  The  objection  to  high  discharge  is  that  the  ore  is 
largely  slimed,  so  that  it  is  difficult  to  concentrate,  and  the  tailing 
loss  in  gold-bearing  pyrite  is  high.  In  California  practice  4  tons 
may  be  crushed  daily;  in  Colorado  practice  but  i  to  1.5  tons. 
\Ye  thus  have  as  an  average  performance  of  a  single  stamp  head : 

California  Colorado 

Drops  per  minute 95  28 

Height  of  drop 7  in.  18  in. 

Height  of  discharge 5  in.  15  in. 

Capacity  in  24  hours 4  tons  1 .25  tons 

Weight  of  stamp i  ,000  Ib.  700  Ib. 

Actual  h.p.  needed  per  stamp      2.02  1.07 

The  theoretical  horsepower  of  a  single  stamp  is  calculated  by 
multiplying  its  weight  into  the  distance  it  is  lifted  per  minute 
divided  by  33,000.  The  actual  horsepower  is  1.2  times  the  the- 
oretical. 

Mercury  fed  to  the  battery. — This  varies  from  i  ounce  to  6 
ounces  per  ounce  of  gold  caught,  the  average  being  about  1.5  oz. 
This  is  added  a  little  at  a  time  inside  the  mortar  and  works  out, 
in  part,  upon  the  apron-plates.  As  to  the  amount  to  use,  a  safe 
guide  is  the  appearance  of  the  plates.  If  the  plates  are  hard  it 
indicates  insufficient  mercury :  if  the  mercury  is  distinctly  visible 
on  the  plates,  either  in  patches  or  streaks,  then  it  is  being  added 


OF    THE    COMMON    METALS.  IO9 

too  freely.  The  mercury  should  be  free  from  base  metals  which 
cause  it  to  'sicken,'  that  is,  it  breaks  up  into  minute  globules  when 
coated  with  these  metals.  Such  globules  refuse  to  coalesce  and 
are  swept  away  in  the  tailing.  Mercury  acts  best  when  it  already 
contains  some  gold  and  silver. 

The  outside  Or  apron-plates  are  dressed  three  or  four  times 
daily,  taking  about  15  minutes  each  time.     To  do  this,  feeding 
is  stopped  to  permit  ore  to  work  out  of  the  battery,  the  stamps 
are  then  hung  up,  and  the  surface  of  the  plate  hosed  off  clean. 
A    rubber-edged    scraper    resembling     a     window    cleaner,    but 
heavier,  is  used  to  scrape  the  plate.     With  this  the  amalgam  is 
gathered  together  and  removed  to  a  suitable  vessel.     An  enam- 
elled kettle  is  commonly  used  for  this 
purpose.     If  the  surface  is  too  hard 
for  the  scraper,  the  amalgam  is  soft- 
ened by  sprinkling  on  a  little  mercury. 
The  plates  are  liable  to  become  tar- 
nished with  the  salts  of  copper  form- 
ing a  verdigris  upon  the  plate  itself ; 
and  since  the  tarnished  part  catches 
no  gold,  such  stains  must  be  removed. 
To  do  this  the  battery  is  stopped,  the 
plate  rinsed  with  clean  water,  and  a 
solution    of    sal-ammoniac    applied    to         FIG.  50.    MERCURY  TRAP. 
the    stained    parts    with    a    scrubbing 

brush.  In  a  few  minutes  this  is  washed  off,  potassium  cyanide 
and  some>  mercury  rubbed  on,  and  the  plate  immediately  washed 
clean. 

Apron-plates  have  a  grade  or  inclination  of  0.5  to  1.75  in.  per 
foot,  heavier  ore  and  sulphides  requiring  the  steepest  grade. 
When  the  pulp  is  flowing  over  the  plate  in  a  proper  manner  it 
travels  down  in  a  series  of  ripples,  its  upper  surface  tumbling 
over  and  over  upon  the  plate,  thus  bringing  the  particles  of  gold 
in  contact  with  it. 

To  save  escaping  particles  of  amalgam  or  of  mercury  which 
fail  to  attach  themselves  to  the  plate,  a  mercury  trap,  Fig.  50,  is 
provided,  especially  where  no  table-concentration  of  the  tailing  is 
attempted.  It  is  in  shape  an  inverted  frustum  of  a  pyramid, 


IIO  THE    METALLURQY 

the  pulp  flowing  in  by  the  vertical  pipe  and  escaping  over  a 
wooden  block  attached  to  the  side  of  the  trap. 

The  clean-up. — Every  two  to  four  weeks  it  is  customary  to 
remove  thoroughly  all  accumulations  of  amalgam  from  the  bat- 
tery. To  do  so  the  stamps  are  hung  up,  two  batteries  at  a  time. 
The  screen,  inside  plates,  and  dies  are  taken  out  and  the  con- 
tents of  the  mortar,  two  or  three  buckets  full,  are  carefully 
scraped  out  and  fed  to  one  of  the  other  batteries.  Finally,  the 
cleanings  of  the  last  batteries  hung  up,  are  put  in  a  clean-up 
pan  (Fig.  51),  together  with  the  amalgam  from  the  well-scraped 
plates.  Three  men  can  clean  40  stamps  in  five  to  seven  hours. 

The  iron  clean-up  pan,  Fig.  51,  3  ft.  diam.  by  2  ft.  deep, 
making  12  to  15  rev.  per  minute,  is  used  for  grinding  the  amal- 
gam, sand,  pyrite,  fragments  of  iron,  and  other  substances,  the 
result  of  the  clean-up.  The  charge,  say  of  300  Ib.  mixed  with 
wrater,  is  ground  for  three  or  four  hours,  after  which  50  Ib.  of 
mercury  is  added  and  the  mixing  is  carried  on  for  a  few  hours 
longer;  the  pulp  is  then  diluted  with  water,  settled  and  dis- 
charged. The  residual  mercury  and  amalgam  is  withdrawn  through 
a  plug-hole  and  strained  through  chamois-skin,  or  through  a  canvas 
bag,  to  remove  the  excess  of  mercury.  Gold  amalgam  when  well 
squeezed  through  the  cloth  will  contain  as  much  as  35  to  45% 
gold.  The  filtered  mercury  still  retains  upwards  -of  0.5%. 

Retorting. — In  the  smaller  gold  mills  the  amalgam  is  retorted 
in  a  pot-shaped  retort  (Fig.  52),  and  in  larger  ones,  in 
a  horizontal  cylindrical  retort  (see  Fig.  53).  The  retort, 
filled  two-thirds  full  of  amalgam,  is  placed  in  an  assay  furnace 
fire  with  the  tube  dipping  into  water,  and  the  charge  is  heated 
slowly  until  the  mercury  begins  to  come  over  and  collect  in  the 
vessel  containing  the  water.  At  this  point  the  fire  is  checked,  and 
the  retort  kept  an  even  heat  for  one  or  two  hours,  after  which 
it  is  brought  to  a  red  heat  to  expel  the  last  of  the  mercury. 

The  mercury,  collected  by  condensation  in  the  water,  is  used 
over  again.  This  accounts  for  its  small  loss,  amounting,  in 
California  practice,  on  an  average  to  0.5  oz.  per  ton  of  ore  treated. 
Mercury  is  lost  both  by  'flouring'  and  by  'sickening' ;  the  first,  a 
white  appearance,  is  due  to  excessive  agitation  and  to  the  air 


OF    THE    COMMON    METALS. 


Ill 


acting  (yi  the  pulp,  while  the  sickened  mercury  is  black-looking 
owing  to  the  presence  of  base  metals,  as  already  explained. 

The  retorted  residue,  still  retaining  0.5  to  i%   Hg.,  is  porous 
and  from  500  to  950  fine  in  gold.     It  is  melted  in  a  plumbago 


FIG.  51.     CLEAN-UP   PAN. 

crucible  with  the  addition  of  some  soda  and  borax,  and  when 
containing  base  metal,  with  a  little  nitre.  The  melt  is  poured  into 
an  ingot  mold,  and  when  cold  the  ingot,  cleaned  from  adherent 
slag,  is  shipped  to  the  mint. 


112  THE    METALLURGY 

When  the  ore  contains  sulphides  which  it  would  pay  to  treat 
by  concentrating,  amalgamation  is  followed  by  concentration  of 
the  tailing  on  tables  like  the  Frue  vanner  or  Wilfley  table.  Thus, 
not  only  is  a  valuable  product  gotten  into  a  small  bulk,  but 
particles  of  amalgam  are  caught  and  also  go  into  the  concen- 
trate. 

When  ores  contain  gold,  both  coarse  and  fine,  the  method  fol- 
lowed has  been  to  amalgamate  the  ore  to  obtain  the  coarse  gold 
and  to  treat  the  tailing  by  the  cyanide  process  to  extract  the  values 
not  obtained  by  amalgamation.  For  further  particulars  of  thus 
treating  gold  ores  see  the  description,  Section  28.  In  South  Africa, 
where  this  method  of  working  obtains,  the  cost  of  milling  may 
be  placed  at  $0.72  to  $1.20  per  ton  for  milling  and  amalgamation, 
and  at  $0.96  to  $1.44  per  ton  for  cyaniding  the  tailing,  all  based 
upon  an  output  of  4.5  to  5.0  tons  per  stamp. 

Cost  of  gold  milling. 

At  a  3O-stamp  mill  in  California  in  1896  there  were  crushed 
and  concentrated  33.512  tons  of  ore  at  an  expense  detailed  as 
follows : — 

Shoes  and  dies $0.029 

Screens 0.003 

Mercury 0.007 

Hardware,  belting,  and  firewood 0.021 

Water  for  power 0.095 

Freight,  cyanide,  oil,  and  grease 0.006 

Lumber 0.008 

Miscellaneous 0.007 

Assay  and  office  supplies 0.008 

Silver-plated  plates 0.007 

Water  pipes  and  connections 0.021 

Hauling  sulphides 0.020 

Express  on  bullion 0.006 

Taxes  and  insurance o.oio 

Superintendence  and  labor 0.160 

$0.408 


OF   THE    COMMON    METALS. 


THE    METALLURGY 


A  summary  shows,  that  of  this  cost,  $0.153  was  for  supplies, 
$0.160  for  labor,  and  $0.095  was  f°r  power.  This  is  for  a  free- 
milling  gold  ore. 

In  Gilpin  county,  Colo.,  where  the  tonnage  put  through  is 
small,  the  cost  was  /8c.  per  ton  in  1891. 

1 8.     LEACH i XG  METHODS  FOR  THE  EXTRACTION  OF  GOLD 
FROM  ORES. 

Gold  occurs  in  ore  in  sizes  from  grains,  or  even  nuggets, 
down  to  particles  of  microscopic  size.  When  in  particles,  in 
sizes  which  show  as  colors  in  panning,  it  is  called  coarse  gold, 
and  such  particles  are  obtained  from  gold  by  amalgamation. 
Much  gold  occurs,  however,  not  only  in  microscopic  particles, 


fefc^ 


2= 


3 i* 


FIG.  53.     HORIZONTAL  RETORT. 

but  also  in  films  upon  the  surfaces  of  pyrite  crystals.  When  in 
this  state,  it  is  capable  of  being  dissolved  by  aqueous  solutions 
containing  chlorine  or  cyanide  of  potassium.  Advantage  is 
taken  of  this  to  get  the  gold  in  solution  by  leaching  or  perco- 
lation methods,  in  which  the  ore  is  treated  in  vats  or  tanks  by 
such  solutions,  the  gold  being  afterward  precipitated  from  the 
clear  filtrate  in  a  flaky  form.  These  flakes  or  particles  are 
caught  on  a  filter,  and  later  dried  and  melted  down  into  gold 
ingots. 

There  are,  therefore,  three  stages  in  obtaining  the  gold  by 
leaching:    First,  the  ore  is  finely  ground,  and,  in  certain  cases, 


OF    THE    COMMON    METALS.  115 

roasted  to  render  the  gold  more  soluble  and  accessible  to  the 
solution ;  second,  it  is  leached  out  by  means  of  a  dilute  solvent 
in  a  tank  having  a  filter-bottom ;  third,  the  gold  is  precipitated 
by  means  of  hydrogen  sulphide  or  by  zinc  from  the  filtrate, 
collected  on  a  filter,  dried  and  melted. 

Up  to  the  present  time  there  are  two  methods  by  which 
gold  may  be  dissolved  from  its  ores,  and  got  into  an  aqueous 
solution  to  be  subsequently  precipitated  as  metal;  viz.,  the 
Plattner  process  in  which  the  gold  is  gotten  into  solution  as 
chloride,  and  the  McArthur-Forrest  process  by  which  it  is  ob- 
tained as  potassium  or  sodium  auro-cyanide. 

19.      CH  LORI  NATION  OF  GOLD  ORES. 

This  process  depends  upon  the  action  of  chlorine  gas  upon 
gold.  The  auric-chloride  is  leached  out  with  water  and  the 
metallic  gold  precipitated  from  the  solution  of  its  chloride  by 
hydrogen  sulphide  or  by  ferrous  sulphate. 

The  ores  best  suited  to  chlorination  are  those  which  cannot 
be  successfully  amalgamated,  such  as  concentrate,  in  which  the 
gold  is  in  a  state  of  fine  division,  and  which  contains  but  little 
silver.  The  latter  as  an  insoluble  chloride  is  apt  to  coat  the 
particles  of  gold.  Since  chlorine  attacks  arsenides  and  anti- 
monides,  ore  containing  them  should  be  subjected  to  an  oxidiz- 
ing roast  to  drive  off  arsenic,  antimony  and  sulphur,  and  to 
oxidize  the  metallic  bases.  A  little  salt  is  added  to  this  roast 
to  chloridize  whatever  metals  are  capable  of  being  so  acted 
upon.  Besides  this,  roasting  makes  the  ore  more  porous,  and 
therefore  more  accessible  to  the  action  of  the  chlorine. 

Roasting  the  ore.  The  ore,  dry-crushed  to  from  10  to  30- 
mesh  size  (see  Section  8),  must  first  be  roasted.  This  is  done 
frequently  in  long  hand-rabbled  reverberatories,  and  also  in 
one  of  the  automatic  roasting  furnaces  such  as  the  Wethey- 
Holthoff  furnace  (see  Fig.  36).  (This  latter  has  a  cooling- 
hearth  where  the  temperature  of  the  ore  may  be  brought  down 
to  a  point  where  it  can  be  charged  to  the  chlorinating  appara- 
tus.) The  roasting  should  be  conducted  at  a  low  temperature, 
which  is  only  raised  at  the  finish  in  order  to  decompose  sul- 
phates as  far  as  possible. 


Il6  THE    METALLURGY 

There  are  two  methods  of  chlorinating  ores,  rat  or  Plattner 
and  the  barrel  or  Thiess  process  of  Chlorination. 

In  the  vat  method  the  ore  is  charged,  in  moistened  condition, 
into  a  tank  or  vat,  and  there  treated  by  chlorine  evolved  in  a 
separate  vessel.  In  barrel  chlorination  the  ore  is  acted  on  in 
rotating  barrels  or  cylinders  by  means  of  chlorine  generated 
within  the  barrel  itself.  The  continuous  movement  of  the  ma- 
terials of  the  charge  brings  the  gold  in  the  ore  intimately  in 
contact  with  the  chlorine,  which  in  nascent  condition  is  sup- 
posed to  act  powerfully  in  promoting  solution. 

Yat  chlorination  recommends  itself  because  of  the  small  in- 
vestment in  plant,  aside  from  the  roasting  furnace  used.  While 
the  capacity  of  such  a  plant  is  small  it  is  sufficient  for  the 
amount  of  concentrate  which  it  is  called  upon  to  treat. 

20.      THE  YAT  OR  PLATTXER  PROCESS  OF  CHLORIXATIOX. 

The  ore,  having  been  subjected  to  an  oxidizing  roast  to  free 
the  gold  and  to  make  the  ore  of  an  open  porous  texture,  is 
moistened  and  charged  into  a  vat  or  tank,  8  or  9  ft.  diam.  by 
3  to  3.5  ft.  deep.  These  vats  (see  Fig.  54)  have  a  false  bottom 
of  perforated  i-in.  boards  supported  on  i-in.  strips.  Upon  the 
perforated  bottom  is  spread  a  6-in.  layer  of  coarse  gravel  of 
hen's  egg  size  at  first,  becoming  smaller  at  the  top.  Above  this 
is  a  layer  of  2  in.  of  sand,  on  which  is  laid  either  a  canvas  filter 
cloth,  or  a  layer  of  boards.  Chlorine  gas  is  admitted  at  a,  and 
drainage  of  the  vat  takes  place  through  the  hose  b  into  the 
launder  c  leading  to  the  settling  vat. 

The  4-ton  charge  of  ore  must  be  moistened  to  the  right  de- 
gree (about  6%),  since,  when  too  dry,  it  is  not  well  acted  on 
by  the  chlorine,  and,  if  too  moist,  the  gas  does  not  pass  through 
it  well.  The  charge  is  thrown  into  the  vat  through  a  screen 
of  o.5-in.  mesh,  which  at  once  breaks  up  all  lumps  and  causes 
the  ore  to  fall  loosely  and  scatteringlv.  When  the  vat  has  been 
filled  to  a  foot  in  depth,  chlorine  is  allowed  to  pass  into  the  space 
below  the  false  bottom  and  rises  through  the  ore.  The  charg- 
ing being  continued,  the  vat  is  filled  to  within  3  in.  of  the  top, 
and  is  covered  with  sacking.  The  cast-iron  cover  d  is  then 


OF    THE    COMMON    METALS.  117 

brought  by  an  overhead  crawl  to  the  tank  and  lowered  upon  h\ 
the  joint  between  being  made  up  of  a  clay-mortar  interlaid 
with  a  cloth  strip  to  keep  it  moist  and  tight.  Chlorine  is  al- 
lowed to  enter  the  vat  for  5  to  12  hours,  according  to  the  fine- 
ness of  the  gold ;  the  finer  the  gold  the  faster  it  is  chlorinated. 
The  charge  is  sufficiently  saturated  with  chlorine  when,  upon 
opening  a  stop-cock  on  the  cover,  escaping  fumes  of  chlorine 


FIG.  54.      CHLORINATION  LEACHING  TANK. 

can  be  detected.  After  this  the  covered  vat  is  allowed  to 
stand  24  to  40  hours  to  chlorinate. 

We  have  at  first : 

Au+3Cl=Cl-(-AuCl2,  which  is  an  insoluble  compound. 
In  presence  of  water,  however,  we  have  the  reaction : 

Aq-|-AuCl2+Cl=Aq-|-AuCl3.  This  latter  salt  is  soluble  in 
water,  so  that  the  gold  can  be  leached.  Hence  the  importance  of 
at  first  moistening  the  ore. 

The  chlorine  is  produced  in  a  generator  (Fig.  55).  This  is  of 
cast  iron,  lead  lined,  and  has  a  heavy,  tight  cover.  It  is  24  in. 
diam.  by  16  in.  deep  for  a  4-ton  charge  of  ore.  To  charge  it  a 


Il8  THE    METALLURGY 

plug  in  the  cover  is  removed  and  the  solid  chemicals  put  in. 
These  are  20  to  27  Ib.  of  binoxide  of  manganese  and  27  to  32  Ib. 
of  common  salt.  From  40  to  60  Ib.  of  sulphuric  acid  of  66°  Be 
is  added  through  the  fissle  tube  u,  followed  by  24  to  33  Ib.  of 
water.  The  chlorine  is  generated  according  to  the  reaction : 

2  NaCl+MnO2+2  H2SO4=2  Cl+Xa2SO4+MnSO4+2  H2O. 

The  apparatus  stands  upon  a  water  or  steam-bath  on  which 
it  can  be  heated  to  50°  C  which  is  the  best  temperature  for  gen- 
erating chlorine.  Through  the  tube  b  the  gas  passes  to  a  wash- 
bottle  to  remove  HC1,  and  thence  to  the  chlorinating  tank. 

The  charge  having  remained  in  the  vat  for  48  hours  is  ready 
for  leaching.  To  do  this,  the  cover  is  removed,  and  the  water 
run  on  over  the  sacking,  it  being  thus  uniformly  distributed  over 
the  ore.  As  soon  as  the  tank  is  quite  full  of  water,  which  no 
longer  settles  through  it,  the  solution  is  allowed  to  escape  through 
the  hose  b,  let  down  for  that  purpose,  the  supply  of  water  being 
kept  up  at  the  top  until  the  escaping  solution  shows  no  reaction 
for  gold.  Two  tons  of  water  is  needed  per  ton  of  ore.  The 
tailing  often  contains  silver  which,  when  in  sufficient  quantity, 
can  be  later  recovered  by  hyposulphite  lixiviation  ( see  Section  45 ) . 

Passing  to  the  settling  tank  the  solution  is  allowed  to  settle 
out  its  sediment  for  several  hours,  after  which  it  goes  to  the 
wooden  precipitating  tank  6  ft.  diam.  by  3  ft.  high,  painted  with 
hot  asphalt  to  which  has  been  added  some  Portland  cement. 
The  precipitant  for  the  gold  is  a  solution  of  ferrous  sulphate, 
prepared  at  the  works  by  dissolving  scrap-iron  in  sulphuric  acid. 
It  precipitates  the  gold  as  follows : 

2  AuCL+6  FeSO4=Au2+Fe2Cl6+2  Fe2(SO4)3 

The  precipitant  is  stirred  in,  after  which  the  vat  is  covered  and 
the  gold  allowed  to  settle  for  at  least  12  hours,  often  much 
longer.  The  supernatant  solution  is  then  run  off,  preferably  to 
a  filter  press,  the  filtrate  from  which  is  later  received  into  a  saw- 
dust filter  to  catch  any  further  particles  of  gold  that  may  have 
escaped  pressing.  Precipitation  with  ferrous  sulphate  has  the 
disadvantage  of  being  very  slow  of  completion,  the  solution 
showing  a  purple  color,  due  to  the  presence  of  gold,  for  days 


OF  THE   COMMON    METALS. 


119 


after  the  precipitation.     It  is  inferior  in  its  action  to  hydrogen 
sulphide. 

The  precipitated  gold  is  filter-pressed  and  washed  with  dilute 
sulphuric  acid,  then  with  hot  water  to  remove  the  ferric  salt  still 
retained  by  it.  It  is  finally  melted  with  soda,  borax  and  nitre 
in  graphite  crucibles.  The  gold  thus  obtained  is  920  to  990  fine, 
the  impurities  being  iron  and  lead.  The  extraction  or  recovery 
of  gold  varies  from  90  to  92%. 
Costs. 

The  cost  of  erecting  a  plant  in  California,  capable  of  treating 
6  tons  daily,  may  be  given  at  $6,000  to  $7,000.     In   1886  the 


FIG.  55      CHLORINE  GENERATOR. 

cost  of  chlorination  at  the  Providence  mine  was  $6.30  per  ton, 
not  ..including  the  items  of  supervision,  interest  and  depreciation, 
about  one-half  being  the  cost  of  roasting.  This  cost  has  since 
been  considerably  reduced  by  the  use  of  fuel  oil  and  automatic 
furnaces  of  the  Edwards  type. 

21.     BARREL  CHLORINATION. 

The  chlorination  barrel. — These  barrels  or  cylinders,  as  shown 
in  Fig.  56,  are  rotated  on  a  horizontal  axis,  being  carried  on 
trunnions,  and  driven  by  spur  gearing  at  12  rev.  per  min.  The 
cylinder  is  lined  with  sheet  lead  bolted  to  the  shell,  and,  as 


I2O  THE    METALLURGY 

shown,  carries  within  it  a  filter-frame  or  diaphragm  of  hard 
wood  intended  to  filter  the  solution  after  treatment,  so  as  to 
obtain  a  filtrate  containing  the  gold  in  solution.  On  this  rests 
lead  plates  0.375  m-  thick,  perforated  with  0.375  m-  holes  0.75  in. 
apart.  The  plate  itself  is  corrugated  for  the  freer  circulation  of 
the  filtrate  over  it.  Instead  of  these  plates  a  board  or  plank 
floor,  similarly  perforated,  has  been  used.  Upon  the  plates 
rests  a  filter  of  sheet  lead  of  a  weight  of  4  Ib.  per  square  foot. 
This  sheet  is  perforated  with  holes  0.05  in.  diam.,  0.375  in.  apart. 
To  hold  down  the  filter-sheet  another  frame  is  fitted  over  it  and 
securely  bolted  down.  The  wood  underframes  last  three  months, 
the  upper  ones  but  two  or  three  weeks.  Barrels  have  been  made 
up  to  1 8  tons  capacity,  6.5  ft.  diam.  by  16  ft.  long.  The  com- 
mon size  is,  however,  5.5  ft.  diam.  by  12  ft.  long. 

Charging. — In  operation,  the  ore,  already  crushed  and  roasted, 
is  drawn  off  from  the  storage  bins  in  two-wheeled  buggies,  and 
is  placed  in  weighed  charges  in  the  charging-hoppers  belonging 
to  each  cylinder.  Into  the  cylinder  is  first  run  135  to  140  gal- 
lons of  water  per  ton  of  ore.  Next  a  measured  quantity  of  sul- 
phuric acid  is  added,  and  upon  this,  the  charge  of  ore.  Lastly  is 
added  a  weighed  amount  of  bleaching  powder.  The  quantity  of 
chemicals  to  be  added  to  the  charge  depends  upon  the  nature  of 
the  ore,  and  is  determined  by  experiment.  On  roasted  Cripple 
Creek  ores  is  used  12  to  15  Ib.  of  bleaching  powder  per  ton  of 
ore  of  34  to  36^  available  chlorine,  and  24  to  30  Ib.  sulphuric 
acid  of  66°  Beaume. 

The  charge-openings  of  the  barrel  are  now  closed,  and  the 
barrel  set  slowly  revolving  (12  rev.  per  min.)  for  a  period  indi- 
cated by  experience,  being  from  one  to  four  hours.  (In  the 'case 
of  Cripple  Creek  ores,  3  hours.) 

The  chemicals  are  brought  thoroughly  in  contact  with  the  ore, 
and  with  one  another,  producing  the  following  reaction : 

2  CaOCl2+2  H2SO4=4Cl+2  CaSO4+2  H2O. 

The  chlorine,  in  formation  in  a  nascent  condition,  acts  with 
greater  energy  upon  the  gold,  forming  a  soluble  gold  chloride 
thus: 

Au+3Cl+H2O=AuCl3H20. 


122  THE    METALLURGY 

To  determine  when  the  ore  has  been  thoroughly  saturated  by- 
chlorine,  a  small  stop-cock  on  the  barrel  is  opened,  and  the  issuing 
gas  tested  for  chlorine  with  ammonia  with  the  formation  of  a 
white  NH4C1  fume.  Another  way  of  furnishing  chlorine,  which 
has  been  used,  is  to  add  it  to  the  barrel  in  liquid  form,  it  being 
obtainable  in  the  market  in  that  form  stored  in  drums. 

The  barrel  is  revolved  for  another  hour,  after  which  it  is 
stopped  with  the  filtering-diaphragm  down ;  the  outlet  pipe  is 
opened  and  connected  by  hose  to  the  settling  tanks,  and  water 
is  pumped  into  the  barrel  above  the  charge.  The  barrel  is  also 
connected  to  the  compressed-air  tank,  and  the  filtering  proceeds 
under  pressure.  The  excess  of  chlorine  is  absorbed  by  the 
water,  and  does  not  enter  the  building.  At  definite  intervals 
leaching  is  suspended,  and  the  barrel  revolved  a  few  times  to 
mix  its  contents  and  to  break  up  channels  which  may  have 
formed  during  the  leaching.  The  compressed  air  is  admitted  at  a 
pressure  of  30  to  40  Ib.  per  square  inch,  and  the  time  of  washing 
and  leaching  is,  on  an  average,  2.5  hours,  the  water  used  being 
50%  by  weight  of  the  ore.  After  the  leaching  is  over  the  barrel 
is  emptied  by  opening  the  man-holes  and  revolving  the  cylinder, 
after  which  the  filter  is  washed  with  a  hose  in  readiness  for 
another  charge. 

The  solution  from  the  barrels  is  received  into  lead-lined  set- 
tling tanks,  10  ft.  diam.  by  7^  ft.  high.  Here  any  sediment  is 
settled  out  during  8  hours,  while  the  supernatant  clear  solution 
is  withdrawn  from  a  point  10  in.  above  the  bottom,  so  as  not 
to  disturb  the  settling,  and  is  sent  to  the  lead-lined  precipitation 
tank  (Fig.  57)  10  ft.  diam.  by  12  ft.  high  entering  it  through  A. 

Precipitation  of  the  gold. — Referring  to  Fig.  57,  precipitation 
of  the  gold  is  performed  as  follows :  The  free  chlorine,  contained 
in  the  gold  solution  in  the  precipitation  tank,  is  first  removed 
by  passing  in  sulphur  dioxide  from  the  generator  //.  This  gen- 
erator contains  a  pan  in  which  lump  sulphur  is  set  on  fire.  Com- 
pressed air,  being  admitted  by  means  of  the  pipe  n,  serves  to  burn 
the  sulphur  and  to  drive  the  fumes  into  the  solution  by  the  pipes 
o  and  u  to  the  lead  pipe  r,  perforated  where  it  crosses  the  bottom 
of  the  tank.  The  SO2  gas  acts  according  to  the  following 
equation:  Cl2+SO,-^2  H2O=H2SO4+2  HC1. 


OF   THE    COMMON    METALS. 


I23 


The  chlorine  having  been  removed,  hydrogen  sulphide  is  passed 
into  the  solution  to  precipitate  the  gold.  Of  late,  treatment  with 
SO2  gas  has  been  omitted,  H2S  gas  being  alone  used.  At  first 
it  is  oxidized  by  the  chlorine,  after  which  it  begins  to  precipitate 


v- 


FIG.  57.     PRECIPITATION  APPARATUS  FOR  BARREL  CHLORINATION. 

the  gold.     Hence,  we  have,  after  removal  of  the  chlorine,  the 
following  reaction: 

2  AuCl3+3  H3S=Au2S3+6  HC1. 

The  precipitation  is  rapid,  taking  only  10  minutes,  and  as 
the  gold  is  thrown  down  before  the  copper,  it  is  possible,  by 
careful  working,  to  leave  the  greater  part  of  the  copper  in 


124  THE    METALLURGY 

solution.  Referring  again  to  Fig.  57,  g  is  the  lead-lined  H2S 
generator.  It  contains  a  perforated  sheet-lead  shelf  on  which 
lies  i -in.  pieces  of  iron  sulphide  upon  which  is  brought  the  tLSO., 
for  the  generation  of  the  H2S.  The  valve  d  having  been  opened 
and  b  closed,  the  gas  passes  through  the  pipes  v  and  r  to  the 
solution,  being  driven  through  it  by  compressed  air  admitted 
to  the  generator  at  c.  When  the  chemicals  in  the  generator  are 
exhausted  they  are  drawn  off  into  a  waste  launder  />.  The  chem- 
icals for  precipitation  needed,  per  ton  of  ore  roasted,  are,  iron 
sulphide  I  lb.,  sulphur  0.25  lb.,  and  sulphuric  acid  2.5  pounds. 

After  precipitation  the  Au2S3  is  allowed  to  settle  for  two 
hours,  after  which  the  supernatant  solution  is  drawn  off  at  C, 
10  in.  above  the  bottom  of  the  tank,  through  the  pipe  u  into  the 
filter  press,  being  under  its  own  hydrostatic  head  of  25  ft.  The 
filtration  is  performed  to  recover  possible  flakes  of  gold  sulphide 
which  may  not  have  settled  to  the  bottom  of  the  tank.  In  3  to  4 
hours  after  precipitation  the  tank  is  ready  to  receive  a  fresh 
charge.  The  life  of  a  set  of  filter-frames  is  equal  to  from  6,000 
to  7,000  tons  of  ore  treated.  The  precipitated  gold  sulphide 
collects  upon  the  bottom  of  the  tank  and,  every  month  or  two 
it  is  drawn  off  through  D,  by  the  rubber  pipe  v,  into  the  pressure 
tank  z,  the  tank  being  also  washed  clean  with  a  hose.  The 
pressure-tank  z  is  4  ft.  diam.  by  4.5  ft.  high.  Into  it  air  is 
forced  through  /  and  its  contents  driven  by  the  pipe  u  through 
the  filter-press  T,  elsewhere  described  and  illustrated  in  Fig.  62. 
The  filtrate  from  the  press,  as  a  safeguard,  passes  over  a  filter- 
bed  placed  in  a  shallow  tank  before  finally  running  to  waste. 

The  precipitate  of  gold  sulphide  contains  a  certain  amount  of 
precipitated  sulphur  together  with  the  sulphides  of  arsenic,  anti- 
mony, copper  and  silver,  forming  a  'sulphide  cake.' 

This  cake  is  mixed  in  trays  44  in.  long,  24  in.  wide,  and  4  in. 
deep,  with  some  borax,  soda  and  nitre,  and  the  trays  are  placed 
in  cast-iron  muffles  heated  by  coal.  These  muffles  are  connected 
with  a  flue-chamber  where  any  mechanically  escaping  dust  may 
be  caught.  In  the  muffles  the  precipitate  is  dried  out,  after  which 
the  heat  is  raised  to  roast  off  and  decompose  the  sulphides,  the 
operation  taking  an  hour.  The  roasted  material  will  have  a  light 
brown  color,  and  contains  upward  of  70  to  80%  An. 


OF    THE    COMMON    METALS.  125 

It  is  now  carefully  transferred  to  a  crucible  and  melted  down 
in  a  wind  furnace.  The  contents  of  the  crucible  are  poured, 
slag  and  all,  into  a  conical  mold,  the  gold  settling  to  the  bottom. 
Upon  cooling,  the  gold  ingot,  900  to  950  fine,  is  separated  from 
the  slag,  re-melted  and  cast  into  a  bar  in  readiness  to  ship  away 
to  the  U.  S.  Mint.  Here,  after  deducting  a  charge  of  2c.  per 
ounce  for  melting  and  assaying,  the  gold  should  net  $20.65  per 
ounce. 

The  slag  resulting  from  this  melt  may  be  re-melted  with  one- 
seventh  of  its  weight  of  litharge  with  the  addition  of  some  re- 
ducing agent,  and  a  lead  button  obtained.  This  lead  can  be  scor- 
ified, and  finally  cupelled  in  the  muffle,  to  obtain  the  small  addi- 
tional amount  of  gold. 

The  above  is  not  the  only  way  of  precipitating  the  gold — pre- 
cipitation by  means  of  charcoal  having  also  been  used.  In  this 
method  the  filtered  solution  from  the  barrels  is  heated  to  50°  C 
to  precipitate  lime-salts  and  to  drive  off  chlorine.  It  then  goes 
through  a  charcoal  filter  40  in.  diam.  made  as  follows :  The 
false  bottom  of  the  filter  is  a  thick  perforated  lead  plate  on  which 
rests  i  cu.  ft.  coarse  charcoal,  then  8  cu.  ft.  fine  charcoal  and 
finally  3  cu.  ft.  of  coarse  charcoal.  The  coarse  charcoal  has  been 
broken  to  lie  between  20  and  30  mesh,  the  fine  between  30  and 
40  mesh.  Both  grades  are  washed  to  free  them  from  dust.  The 
gold  solution  is  filtered  hot  at  the  rate  of  20  gallons  per  hour 
leaving  the  solution  with  0.003  to  0.007  oz.  gold  only.  The  gold 
shows  as  a  gilded  appearance  upon  the  surfaces  of  the  particles 
of  charcoal,  especially  at  the  top  layers  where  most  of  the  pre- 
cipitation occurs.  The  exhausted  filtrate  is  run  into  a  tank 
having  a  lead-plate  false-bottom  and  containing  sawdust  to  the 
depth  of  a  foot.  Yearly  this  sawdust  is  burned  and  the  ashes 
sold  to  the  smelter. 

The  charcoal  is  periodically  removed  from  the  filter  and  burned 
upon  flat  plates,  from  which  the  ashes  can  be  carefully  swept  up. 
These  ashes,  containing  5  to  10%  of  gold,  are  melted  in  crucibles 
with  their  own  weight  of  a  I  to  I  mixture  of  soda-ash  and  borax. 
The  melt  is  poured  into  a  conical  mold,  and  upon  cooling  the 
gold  is  removed  and  re-melted  into  an  ingot.  The  slag  is  sold 
to  the  smelters. 


126  THE    METALLURGY 

The  extraction  of  gold  is  90  to  94%  according  to  the  nature 
of  the  ore.  The  silver  is  not  recovered,  being  in  the  form  of  an 
insoluble  silver  chloride.  However,  if  some  salt  has  been  used 
in  the  preliminary  roasting,  the  silver  chloride  then  formed,  if 
insufficient  in  quantity  to  pay,  may  be  extracted  by  further  treat- 
ment, leaching  it  by  means  of  sodium  hyposulphite  with  a  re- 
covery of  60%. 

We  give,  herewith  the  cost  per  ton  of  treating  Cripple  Creek 
ores  on  a  large  scale  by  barrel  chlorination : 

Labor,    including    salaries $i-34 

Chemicals  and  supplies 0.72 

Fuel,   roasting   and   power 0.70 

Renewals   and   repairs 0.45 

Miscellaneous   expenses 0.32 


Total  cost  per  ton. $3-53 

The  chemicals  used  in  the  process  are:  Sulphuric  acid  (oil 
of  vitriol  of  66°  Be.)  $0.90  to  $1.10  per  100  Ib. ;  chloride  of  lime 
(bleaching  powder)  $1.80  per  100  Ib.,  (New  York)  ;  sulphide  of 
iron  $3  per  100  Ib.,  or  it  can  be  made  at  the  works  from  wrought- 
iron  scrap  and  sulphur;  sulphur  $2  per  100  Ib. 

The  quantity  of  water  needed  for  high-pressure  steam  power 
and  for  chlorination  will  be  2  tons  per  ton  of  ore  treated.  This 
can  be  reduced  if  settling  tanks  are  employed  and  the  water  used 
over  again. 

22.    CVAXIDIXG  OF  GOLD  (AND  SlLVER)  ORES. GENERAL 

OBSERVATIONS. 

The  process  of  cyaniding  consists  essentially  in  attacking  gold 
and  silver  ores  by  dilute  solutions  containing  less  than  0.5% 
cyanide  of  potassium  (caustic  soda  or  lime  being  added  to  ores 
which  have  been  rendered  acid  by  the  oxidation  of  pyrite)  and 
then  in  precipitating  the  precious  metals  by  means  of  zinc  shav- 
ings or  zinc  dust,  or  by  electrolysis. 

The  process  is  a  success  with  many  ores,  and  its  field  of  use- 
fulness is  bound  to  be  greatly  extended,  partly  at  the  expense 
of  the  chlorination  process. 


OF   THE    COMMON    METALS.  127 

The  chief  advantage  of  cyaniding  over  chlorination  is  that 
roasting  is  by  no  means  always  essential  even  when  sulphides 
are  present.  This  is  an  important  point  in  the  treatment  of 
low-grade  ores,  especially  where  fuel  and  labor  are  costly. 

It  will  be  noticed  that  the  dilute  cyanide  solution  must  have  a 
selective  action,  that  is,  it  must  be  able  to  dissolve  the  gold  and 
silver  without  attacking  the  base  metals  which  may  be  present. 
If,  however,  copper  is  present,  it  will  consume  cyanide,  though  the 
presence  of  copper  in  small  quantities  is  now  no  longer  considered 
a  serious  obstacle  to  cyaniding. 

The  ores  to  which  the  process  is  adapted  are: 

1.  Free-milling  ores  in  which  the  gold  is  in  fine  or  microscopic 
particles.     If  the  gold  is  in  coarse  particles  it  takes  so  long  to 
dissolve  it  that  the  method  would  be  impracticable.     However, 
it  is  possible  to  remove  the  coarse  gold  by  amalgamation,  after 
which  the  residue  can  be  treated  by  cyaniding. 

2.  Telluride  ores,   which  have  first  been  roasted  to  set  the 
gold  free  from  its  combinations  with  tellurium. 

3.  Pyritic  ores  in  which  the  gold  occurs  upon  the  faces  of 
the  crystals  of  pyrite  as  films.    When  this  is  crushed  fine  enough 
to  expose  these  faces  the  gold  is  attacked  by  the  solution.     Some 
pyrite  ores,   however,  have  the  gold  included  in  the   substance 
of  the  crystals  and  hence  may  be  profitably  roasted.    Roasting  has 
also   the   advantage   that   it   renders   the   ore   more   porous   and 
therefore  more  accessible  to  the  solution. 

Talcose  or  clayey  ores,  when  crushed  for  cyaniding,  make  a 
great  deal  of  slime  which  is  exceedingly  difficult,  if  not  impossi- 
ble, to  leach,  and  hence  require  treatment  as  a  slime,  adding  much 
to  the  cost  of  extraction. 

23.     DEVELOPMENT   OF  THE   CYANIDE   PROCESS. 

Beginning  in  South  Africa  in  1889,  the  process  was  at  first 
applied  to  the  treatment  of  impounded  tailing  from  the  gold 
stamp-mill  to  recover  the  gold  not  extracted  by  amalgamation. 
This  tailing  was  leached  in  vats  with  cyanide  solution  to  dissolve 
the  gold  which  was  then  precipitated  by  sending  the  solution 
through  zinc  boxes  full  of  zinc  shavings.  So  long  as  these 


128  THE    METALLURGY 

dumps  lasted,  this  simple  treatment  was  sufficient,  but  when 
finally  the  tailing  had  to  be  treated  as  fast  as  it  was  produced, 
the  practice  was  modified  as  follows : 

The  pulp  as  it  left  the  amalgamation  plates  was  run  through 
classifying  boxes  which  gave  two  products.  The  first  of  these 
was  the  coarse  sand  or  concentrate.  This  was  treated  for  a  long 
time  and  with  comparatively  stronger  solution  to  effect  recovery. 
The  second  product  was  the  mingled  slime  and  sand  whose  acidity 
was  neutralized  by  the  addition  of  a  regulated  amount  of  milk- 
of-lime  which  flowed  into  a  launder  which  led  to  a  Butters  & 
Mein  distributor  (Fig.  77).  This  latter  delivered  it  evenly  to 
a  settling  vat,  the  clean  sand  settling  out  while  the  slime  was 
carried  away  in  the  escaping  turbid  water ;  when  the  settling 
vat  had  been  filled,  its  excess  of  moisture  was  displaced  by  a 
weak  solution,  after  which  the  contents  of  the  vat  were  shoveled 
through  trap-doors  into  a  leaching  vat  set  immediately  under  the 
upper  one.  Strong  solution  containing  0.15  to  0.25%  potassium 
cyanide  was  then  allowed  to  percolate  through  this  sand  for  a 
certain  number  of  days,  was  displaced  by  weak  solution,  and 
finally  by  water.  These  solutions,  as  they  left  the  vat,  were  run 
through  the  zinc  boxes  to  precipitate  their  contained  gold.  The 
exhausted  tailing  was  then  shoveled  out  and  trammed  away. 

The  slime  was  caught  in  large  vats,  allowed  to  settle,  and 
the  supernatant  water  decanted.  A  very  dilute  cyanide  solution 
was  added  to  the  remaining  slime  and  the  whole  was  agitated 
by  a  mechanical  stirrer.  After  some  hours'  treatment  this  mix- 
ture was  allowed  to  settle  and  the  clear  liquid  precipitated  by 
zinc  shavings. 

Since  settling  by  decantation  was  slow,  a  large  number  of 
rats  was  needed  and  moreover  the  last  portion  of  the  gold  could 
not  be  saved,  so  that  decantation  was  succeeded  by  the  filter- 
press  method  of  slime  treatment.  This  consisted  in  filter-press- 
ing the  agitated  slime  under  high  pressure  through  such  a  press 
more  fully  described  in  Section  31. 

The  comparatively  high  cost  of  this  system  led  to  the  adoption 
of  the  Moore  system  and  to  similar  methods  of  suction  filter- 
pressing.  The  most  successful  of  these  is  that  of  Cassel  & 
Butters. 


OF    THE    COMMON    METALS.  I2Q 

The  West  Australian  ores  are  of  such  a  nature  that,  to  liberate 
their  contained  gold,  fine  grinding  was  needed,  and  while  this 
fine  material  could  not  be  leached  in  vats,  it  could  be  filter- 
pressed.  This  led  to  the  use  of  the  grinding  pan  as  employed  in 
silver  amalgamation  and  chiefly  to  the  adoption  of  the  tube-mill, 
which  has  proved  itself  particularly  efficient  in  fine  grinding. 

The  trend  of  modern  practice  is  toward  fine  grinding  and  to 
the  treatment  of  the  whole  product  either  as  a  slime,  or  else  a 
separation  of  the  ground  product  into  sand  and  slime,  the  former 
to  be  treated  by  percolation,  the  latter  by  a  filtering  process. 

24.     CHEMISTRY  OF  THE  CYANIDE  PROCESS. 

When  a  dilute  solution  of  less  than  0.5%  of  potassium  cyanide 
is  brought  in  contact  with  finely  ground  ore,  in  which  gold  occurs 
in  metallic  finely  divided  particles,  the  gold  goes  into  solution 
according  to  the  reaction  first  set  forth  by  Eisner  as  follows : 
2  Au+4KCN+O+H2O=2  AuK  (CN)2+2  KOH. 
or  with  the  formation  of  an  auro-potassium  cyanide  and  of 
caustic  potash.  According  to  this  reaction  i  oz.  of  potassium 
cyanide  should  dissolve  1.5  oz.  of  gold,  but,  in  practice,  30  to  40 
times  this  quantity  is  required.  The  reaction  ceases  when  the 
oxygen  of  the  dissolved  air  is  gone,  and  begins  again  with  the 
access  of  fresh  air.  Oxidizing  agents,  as  potassium  chlorate  and 
permanganate,  and  the  peroxides  of  lead,  manganese  and  sodium 
may  replace  air  in  furnishing  oxygen,  but  are  more  expensive. 
Silver  may  be  dissolved  as  well  as  gold  by  the  same  reaction, 
namely : 

2  Ag+4KCN+O+H2O=2  AgK(CN)2+2  KOH. 

Silver  dissolves  less   readily,   however,  than  gold. 

Another  powerful  reagent  has  been  found  in  bromo-cyanide. 
made  from  the  crystals  or  at  the  works  by  adding  bromine  water 
to  cyanide  solution,  which  reacts  with  potassium  cyanide  thus : 

KCN+BrCN=KBr+C2N2 

forming  cyanogen  (C2N2)  which  in  nascent  condition  acts  directly 
on  the  gold  as  follows : 

2  Au+C2N2+2  KCN=2  K  Au  (CN)2 


I3O  THE    METALLURGY 

Bromo-cyanide,  which  does  not  require  the  aeration  of  the 
solution,  as  when  the  ordinary  cyanide  solution  is  used,  is,  how- 
ever, more  expensive,  and  has  been  applied  only  to  those  sulpho- 
tellurides  which  are  unattacked  or  but  slowly  attacked  by  the 
latter. 

When  pyritic  ore  has  been  mixed  and  left  out  in  the  weather, 
air  and  moisture  act  on  it  thus: 

3   FeS2+2   H2O+22O=Fe   SO4+Fe,(SO4)s+2   H2SO4 

\Yhile  this  action  is  apt  to  be  superficial,  still  the  ore  gives  an 
acid  reaction,  and  decomposes  cyanide  solution  with  the  forma- 
tion of  hydrocyanic  acid.  To  prevent  this  the  ore  should  be  given 
a  water-wash  to  remove  the  sulphuric  acid  and  ferrous  sulphate, 
while  the  ferric  sulphate  may  be  neutralized  by  a  caustic  soda 
solution,  or  preferably,  by  lime-water  which  converts  it  into  a 
harmless  hydrate. 

To  precipitate  gold  from  its  solution,  zinc,  either  in  the  form 
of  shavings  or  as  zinc  dust,  is  used ;  the  gold  being  obtained  ar. 
a  black  or  brown  precipitate  thus : 

2  KAu   (CX.)2+Zn=K2  Zn(CX)4+Aiu 

According  to  the  equation  one  part  zinc  should  precipitate  6.2 
parts  gold,  but  in  actual  practice  one  part  of  zinc  is  needed  for 
0.2  to  0.07  part  of  gold,  much  zinc  being  used  up  to  satisfy  re- 
actions produced  by  other  substances  in  the  solution. 

25.     OPERATION  OF  PLANT. 

The  modes  of  operating  cyanide  plants  will  vary  in  accordance 
with  whether  the  plant  is  treating  dry  crushed  ore,  sand  and  slime, 
or  slime  alone.  In  the  treatment  of  dry  crushed  ore,  and  of  sand 
in  direct-treatment  plants  the  leaching  is  an  important  operation. 

Leaching. — The  fineness  to  which  ore  must  be  crushed  in  order 
to  give  the  best  results  by  leaching  must  be  determined  for  each 
case.  With  the  porous  ore  coarse  crushing  is  permissible.  With 
some  hard  ores  or  those  in  which  the  gold  is  uniformly  dis- 
tributed and  very  fine,  it  is  necessary  to  crush  fine  or  even  slime 
the  material  in  order  to  get  satisfactory  extraction.  "We  will 
assume  an  average  case  of  a  sand  of  which  about  50%  will  pass 


OF  THE  COMMON    METALS.  13! 

A  ioo-mesh  screen.  If  impounded  tailing  or  dry  crushed  ore, 
it  is  at  once  delivered  to  the  leaching  tanks,  varying  in  diameter 
from  12  to  40  feet,  by  means  of  cars  or  belt  conveyors,  and 
evenly  distributed.  If  delivered  to  the  tank  wet  after  crushing 
in  stamps,  it  has  either  had  a  preliminary  settling,  or  the  settling 
and  leaching  may  be  accomplished  in  the  same  tank,  as  at  the 
Homestake  mills  in  South  Dakota.  In  either  case  the  charge  is 
simply  leveled  off,  and  the  strong  solution  run  on  the  top  or 


FIG.  58.     BLAISDELL  EXCAVATOR. 

introduced  from  underneath  the  filter.  The  method  of  intro- 
ducing solution  varies  greatly  in  different  mills.  This  strong 
solution  contains  from  0.2  to  0.3%  potassium  cyanide;  if  the 
ore  contains  much  silver,  the  strong  solution  may  be  as  strong 
as  0.4%  cyanide.  The  proper  strength  of  solution  is  an  im- 
portant consideration  and  must  be  determined  experimentally. 
When  the  charge  is  saturated  it  is  customary  to  allow  it  to  soak 
for  several  hours  in  order  to  permeate  the  charge  thoroughly 
and  thereby  avoid  channeling  during  percolation.  The  valve 


132  THE    METALLURGY 

under  the  filter  is  then  opened  and  the  solution  allowed  to  run 
to  the  gold  solution  tank.  Thence  it  passes  to  the  zinc-boxes, 
to  the  sumps,  and  finally  to  the  storage  tanks  where  it  is  re- 
enforced  with  more  cyanide  and  used  over  again  on  the  charge 
Thus  the  charge  is  subjected  to  a  series  of  percolations  with 
strong-  solution,  the  solution  being  allowed  each  time  to  disappear 
below  the  surface  so  as  to  draw  down  air  into  the  sand  and 
accelerate  the  extraction.  The  strong  solution  may  be  applied 
for  several  days.  With  gold  ores  from  4  to  6  days  are  usually 
sufficient.  With  silver  ores  leaching  from  10  to  20  days  is  re- 
quired. When  the  final  strong  solution  is  drawn  off,  a  weak 
solution  of  from  one-third  to  one-half  the  strength  of  the  strong 
is  applied ;  and  finally  a  wash  of  water  is  run  on  to  wash  out 
the  last  traces  of  dissolved  metals  before  the  residues  are  dis- 
charged. These  residues  may  be  discharged  in  various  ways : 
Into  cars  through  side  doors  (see  Fig.  72)  ;  or  sluiced  out  with 
water  into  a  flume  beneath  the  tanks,  or  removed  by  means  of  a 
very  effective  device  known  as  the  Blaisdell  excavator,  which  is 
now  used  in  large  modern  plants  and  effects  a  great  saving  in 
labor  cost  (Fig.  58). 

Slime  treatment. — In  all  wet  crushing  mills  the  production 
of  a  certain  amount  of  slime  is  unavoidable,  that  is,  a  product 
composed  of  such  finely  divided  particles  of  quartz  and  clay 
that  it  cannot  be  treated  by  percolation.  This  so-called  'slime' 
may  consist  of  from  20  to  40%  of  the  total  ore  crushed.  As  it 
requires  a  separate  treatment,  it  is  usually  segregated  by  means 
of  classifiers  of  which  there  are  a  great  variety,  the  cone  classifier 
with  a  central  intake  and  a  peripheral  overflow  being  now  most 
commonly  used  in  the  United  States  (Fig.  76).  As  we  have 
already  noted,  when  the  cyanide  process  was  first  introduced 
into  South  Africa  the  sand  alone  was  treated,  the  bulk  of  the 
finest  slime  having  been  allowed  to  overflow  from  the  dams. 
But  as  the  slime  carries  about  as  much  metal  as  the  sand,  and 
in  some  instances  considerably  more,  the  necessity  of  treating 
this  product  was  soon  recognized.  The  great  difficulty  besetting 
the  slime  problem  was  to  find  some  cheap  method  of  removing 
the  slime  solids  from  the  solution  carrying  the  precious  metals. 
As  we  have  seen,  decantation  was  at  first  tried  and  is  even  now 


OF   THE    COMMON    METALS.  133 

used  in  South  Africa,  being  the  cheapest  available  method  for 
handling  the  low-grade  slime  in  that  field.  This  was  followed 
by  the  introduction  of  filter  presses  in  West  Australia  and  the 
vacuum  filter  in  the  United  States. 

In  view  of  recent  developments  in  cyanide  practice  a  plant 
for  the  treatment  of  slime  is  a  necessary  adjunct  of  nearly  all 
mills  for  the  wet  reduction  of  gold  and  silver  ores.  The  first 
step  in  the  treatment  is  the  settling  of  the  slime.  To  accom- 
plish this  the  slime  overflow  from  the  classifiers  is  conducted 
to  the  centre  of  a  cone-bottom  settling  tank  (Fig.  76)  and  dropped 
into  a  cylinder  of  about  12-in.  cliam.,  which  is  suspended  over  the 
centre  of  the  tank  and  extends  about  half  way  to  the  bottom. 
By  this  means  the  agitation  of  the  falling  slime  is  confined  to 
the  interior  of  the  cylinder :  while  the  surface  of  water  outside  the 
cylinder  is  perfectly  quiet.  These  settlers  are  of  varying  size. 
In  operation  the  clear  water  rises  around  the  central  cylinder 
while  the  slime  settles  to  the  bottom :  and  the  clear  water  finally 
overflows  at  the  top  of  the  tank  into  a  circular  launder  provided 
for  that  purpose.  This  settling  and  clarifying  operation  goes 
on  for  about  24  hours.  The  number  of  tanks  required  for  this 
purpose  will  obviously  depend  upon  the  capacity  desired  and 
the  rapidity  with  which  the  slime  settles.  The  capacity  of  the 
tank  must  be  carefully  estimated  for  any  particular  slime,  as 
there  is  as  much  difference  in  the  character  of  slime  and  its  rate 
of  settling  as  in  ores  themselves.  It  must  be  borne  in  mind,  too. 
that  the  settled  charge  of  slime  becomes  less  dense  as  the  filling 
proceeds ;  consequently  at  a  certain  time  the  separation  of  water 
from  slime  is  no  longer  perfect  and  the  slime  begins  to  flow  over 
with  the  water.  When  this  occurs,  the  stream  should  be  at  once 
shifted  to  another  tank,  and  the  contents  of  the  first  settler  al- 
lowed to  settle  undisturbed.  Usually  twelve  to  twenty-four  hours 
are  allowed  for  settling.  The  settler  is  provided  with  a  decant- 
ing apparatus  consisting  of  a  pipe  connected  with  a  flange  at 
the  bottom  of  the  tank  by  means  of  a  swivel  joint,  which  permits 
of  raising  and  lowering  the  pipe  at  will.  Unless  the  pulp  has 
already  been  crushed  in  a  cyanide  solution  (see  practice  in  Black 
Hills)  enough  cyanide  solution  is  added  to  the  pulp,  now  con- 
taining about  50^  moisture,  to  raise  the  whole  to  the  desired 


134  THE    METALLURGY 

strength  and  to  make  a  consistency  of  about  3  parts  solution 
to  one  of  dry  slime.  This  pulp  is  then  transferred  by  gravity 
or  by  centrifugal  pump  to  the  agitator.  In  some  plants  the 
same  tank  answers  for  both  settler  and  agitator.  The  choice 
must  be  governed  by  conditions.  In  general,  the  settler  should 
be  large  in  diameter  at  the  expense  of  height;  the  agitator,  just 
the  reverse,  especially  if  the  centrifugal  pump  is  to  be  used  for 
agitating.  Moreover,  the  object  of  the  agitator  is  to  prevent 
settling,  hence  it  is  desirable  that  the  bottom  have  a  steeper  slope 
than  the  settler.  In  very  large  tanks,  however,  the  cone-bottom 
may  be  dispensed  with,  owing  to  difficulties  and  expense  of  con- 
struction :  and  all  such  tanks  may  be  provided  with  stirring  de- 
vices to  facilitate  emptying. 

26.     CLASSIFICATION  OF  METHODS. 

First  method. — Where  the  ore  is  soft  and  oxidized,  and  no 
advantage  is  to  be  gained  by  amalgamation  and  concentration, 
it  is  crushed  dry  in  roils  and  treated  by  leaching,  either  with  or 
without  roasting;  examples,  Cripple  Creek,  Colorado;  Mercur, 
Utah. 

Second  method. — Where  the  ore  is  stamped,  then  treated  by 
amalgamation  or  concentration,  or  both,  and  the  tailing  treated 
by  cyanide  after  separation  of  sand  from  slime.  This  class  may 
be  divided  into  three  sub-classes : 

A.  Mills  in  which  the  slime  is  treated  by  decantation ;  best 
example,  South  Africa. 

B.  Mills  in  which  the  slime  is  treated  by  filter-pressing;  best 
examples,  New  Zealand;  Homestake  mine,  South  Dakota. 

C.  Mills  in  which  the  slime  is  treated  by  the  vacuum  filter; 
best  examples,  Terry,  South  Dakota;  Liberty  Bell  mine,  Colo- 
rado; Goldfield  and  Virginia,  Nevada. 

Third  method. — Where  the  ore  is  crushed  in  stamps  and 
either  a  large  part  or  the  whole  of  it  re-ground  or  slimed  in 
tube-mills.  ( I )  At  El  Oro,  Mexico,  a  large  part  of  the  ore  is  re- 
ground  in  tube-mills;  the  fine  sand  is  leached  and  the  slime 
treated  by  decantation.  (2)  At  the  Liberty  Bell  mill,  Telluride, 
Colorado,  the  whole  product  from  80  stamps  is  reduced  to  a 


OF   THE    COMMON    METALS.  135 

slime  in  tube-mills.  This  product  is  then  treated  by  agitation 
and  filtered  on  the  Moore  vacuum  filter. 

Fourth  method. — Where  the  ore  is  crushed  in  stamps,  amal- 
gamation and  concentration  omitted,  the  sand  leached  and  the 
slime  treated  by  decantation.  The  interesting  feature  of  this 
method  is  that  the  ore  is  crushed  in  the  cyanide  solution.  The 
process  is  applied  successfully  at  large  plants  in  the  Black  Hills, 
South  Dakota. 

Fifth  method. — The  reduction  of  the  ore  to  a  slime,  followed 
by  filter-pressing ;  example,  West  Australia. 

27.    FIRST  METHOD  OF  CYANIDING. 

This  is  well  suited  to  the  treatment  of  ores,  that  do  not  pro- 
duce much  slime  and  do  not  require  amalgamation  and  concen- 
tration;  or  to  ores  which  require  roasting.  It  must  not  be  for- 
gotten that  silver,  as  well  as  gold,  is  extracted  by  cyaniding, 
though  not  to  the  same  extent. 

The  conditions  requiring  dry  crushing  are  comparatively  rare, 
and  in  many  instances,  for  economical  reasons,  dry  crushing 
plants  have  been  converted  into  wet  crushing.  Where  efficient 
rolls  can  be  had,  dry 'crushing  is  preferred  for  the  following 
reasons : 

The  ore  is  delivered  to  the  tank  dry  and  therefore  there  is  no 
dilution  of  the  solution  as  in  treating  wet  ore,  and  in  the  treat- 
ment of  dry-crushed  ore  less  water  is  required. 

A  more  uniform  and  more  easily  percolated  bed  can  be  ob- 
tained and  the  ore  is  better  aerated,  the  oxygen  present  assisting, 
as  we  have  already  seen,  in  the  dissolution  of  the  gold. 

Finally,  dry-crushing  by  rolls  ensures  a  more  granular  product 
than  is  attainable  by  stamp-milling. 

On  the  other  hand,  the  cost  of  dry-crushing  beyond  a  certain 
mesh  is  excessive;  and  except  in  rare  instances,  it  is  not  practi- 
cable to  crush  fine  enough  in  rolls  for  the  highest  extraction. 

The  rate  of  percolation  of  dry-crushed  material  will  vary  in 
accordance  with  the  proportion  of  slimy  material  (talc  or  clay). 
So  long  as  that  rate  is  not  less  than  I  in.  per  hour  such  material 
can  be  treated  without  first  removing  the  slime  by  methods  to 
be  described  later. 


136  THE    METALLURGY 

Fig.  59  and  60  represent,  in  plan  and  sectional  elevation,  a  75- 
ton  plant  designed  to  treat  impounded  tailing  containing  so  little 
slime  that  only  37%  of  the  material  will  pass  an  8o-mesh  screen, 
so  that  a  charge  of  it  may  be  treated  in  four  days,  a  is  the  strong- 
solution  tank  12.5  ft.  diam.  by  10.5  ft.  deep,  containing  a  stock 
solution  of,  for  example  0.25%  potassium  cyanide,  or  5  Ib.  per 
ton.  a'  of  the  same  size,  holds  the  weak  solution,  which  may  be 
specified  as  containing  0.1%  or  2  Ib.  KCX  per  ton.  Both  these 
tanks  are  situated  above,  and  supply  the  leaching-vats,  b  b  b  b, 
each  20  ft.  diam.  by  6.5  ft.  deep  inside,  having  a  capacity  of  75 
tons  of  ore,  or  85  Ib.  per  cu.  ft.  The  strong-solution  stock-tank  is 
connected  to  both  top  and  bottom  of  the  leaching  vats  by  the  pipe, 
C,  while  Cf  supplies  the  vats  at  the  top.  A  pipe,  X,  beneath  the 
charging  platform,  gives  a  supply  of  water  for  sluicing  out  the 
contents  of  the  vats  where,  as  in  this  case,  that  method  is  used  for 
the  removal  of  the  exhautsed  ore.  Y  is  a  platform  placed  4.5  ft. 
below  the  top  of  the  vats  for  convenience  of  access.  Z  is  the 
charging  platform,  not  shown  in  the  plan,  Fig.  60,  by  which  ore  is 
brought  in  and  dumped  into  the  leaching-vats.  There  are  two 
gold-solution  vats,  h  h'  (strong  and  weak  solutions)  12.5  ft.  by 
5.5  ft.  deep,  which  receive  the  filtrate  from  the  leaching  vats  by 
way  of  launders,  d  and  d' ',  and  deliver  them  in  a  regulated  stream 
to  the  zinc-boxes,  the  strong  solutions  passing  through  K  K  K  to 
the  sump-tank  /,  and  the  weak  solutions  through  K'  K'  to  the  tank 
/'.  There  are  five  sets  of  these  boxes,  each  containing  9  com- 
partments 12  in.  deep,  15  in.  long  and  24  in.  wide.  W 
is  a  centrifugal  pump  by  which  the  contents  of  the  sumps 
may  be  returned  to  their  respective  stock-tanks  a  and  a'. 
By  means  of  the  vacuum  tank  5"  and  pump  V  a  vacuum 
may  be  created  beneath  the  false  bottom  of  the  leaching- 
vats,  thus  increasing  the  rate  of  percolation.  The  liquid  thus  re- 
moved is  discharged  from  the  tank  by  the  pipe  U  to  either  of  the 
gold-tanks  from  their  respective  sump-tanks.  On  the  near  side 
of  each  of  the  zinc-boxes  is  to  be  seen  a  double  line  representing 
the  launders,  by  which  gold  precipitate  is  carried  from  any  of 
the  boxes  and  delivered  through  the  launders  n  and  m  to  the 
'acid  tank'  0.  In  this  tank  the  gold  precipitate  is  treated  with 
sulphuric  acid  for  the  removal  of  the  entangled  zinc,  the  clear 


OF   THE    COMMON    METALS. 


137 


138  THE    METALLURGY 

supernatant  liquid  being  decanted  to  the  settling  tank  p,  while 
the  settlings  of  the  acid  tank  go  to  the  filter-press  r.  The  other 
apparatus  consists  of  a  gasoline  engine,  or  a  motor,  which  fur- 
nishes power  to  operate  the  pumps,  a  drying  furnace  for  drying 
the  precipitate  and  a  melting  furnace  where  the  dried  precipitate  is 
reduced  to  bullion. 

A  plant,  such  as  above  described,  costs  $150  to  $200  per  ton 
of  ore  capacity  in  24  hours.  This  does  not,  of  course,  include 
the  plant  for  crushing  the  ore. 

Precipitation. — The  strong  and  weak  solutions,  leaving  the 
leaching  vats,  are  gathered  in  their  respective  gold-tanks,  and 
then  passed  in  a  regulated  flow  from  the  zinc-boxes  into  the 
sumps,  the  gold  in  the  solution  being  precipitated  on  zinc  shav- 
ings contained  therein.  Zinc  shavings  have  the  advantage  of  per- 
mitting the  free  passage  of  solution,  and  do  not  clog  the  screen 
on  which  the  gold  precipitate  falls.  They  may  be  bought  ready 
made,  or  can  be  made  at  the  works,  since,  when  freshly  made, 
they  are  more  efficient.  To  make  them  sheets  of  zinc  are  wound 
around  a  mandrel  in  a  lathe,  the  edge  of  the  sheet  being  then 
soldered  down.  A  side-cutting  tool  is  used  to  cut  the  shavings, 
which  are  about  one  twelve-hundredth  of  an  inch  in  thickness, 
0.03  in.  wide,  and  several  feet  in  length.  All  the  compartments 
of  a  set  of  boxes,  except  the  last,  are  loosely  but  uniformly  filled 
with  the  shavings.  The  strong  solution  in  the  plant  ( Fig.  59  and 
60)  flows  through  the  three  sets  of  boxes  at  the  rate  of  one  ton 
per  hour.  The  gold  is  precipitated  in  the  first  compartments,  so 
that  by  the  time  the  solution  reaches  the  ninth  compartment,  no 
discoloration  of  the  shavings  is  to  be  noticed,  and  the  barren  so- 
lution goes  to  the  sump  L.  The  deposit  on  the  zinc  has  a  brown- 
ish or  grayish  black  hue.  As  it  increases  in  the  first  compart- 
ment, the  shavings  become  soft  and  stringy,  the  precipitate,  and 
finally  the  zinc  in  small  pieces  (short  zinc),  settling  to  the  bot- 
tom of  the  compartment.  As  the  shavings  in  the  first  compart- 
ment settle  down,  it  is  usual  to  replenish  them  with  zinc  from  the 
last  compartments,  which  in  turn  are  filled  with  fresh  shavings. 
Gold  is  precipitated  with  greater  difficulty  from  weak  solutions, 
and  accordingly  it  passes  in  series  through  both  sets  of  boxes  K' , 
K'  to  the  sump  L.  From  95  to  99%  of  the  gold  should  be  pre- 


OF   THE    COMMON    METALS. 


139 


140  THE    METALLURGY 

cipitated  by  the  zinc,  and  since  the  barren  solution  is  pumped 
back  to  the  stock  tanks  and  used  again,  this  residual  gold  is 
not  lost. 

Cleaning  up,  acid  treatment,  and  refining  of  the  precipitate. — 
This  may  be  done  monthly  or  bi-monthly,  according  to  the  rich- 
ness of  the  ore  and  the  need  of  realizing  values,  and  is  con- 
ducted as  follows :  The  flow  of  gold  solution  is  stopped  and  to 
displace  the  solution  water  is  run  through  the  one  set  of  boxes 
which  is  about  to  be  cleaned  up.  Beginning  in  any  compartment, 
and  with  the  hands  protected  by  rubber  gloves,  the  contents  are 
agitated  by  raising  and  lowering  the  zinc  shavings  for  a  few 
minutes,  being  careful  to  do  it  gently,  so  as  to  make  as  little 
'short  zinc'  as  possible.  The  water  becomes  black  \vith  the  float- 
ing precipitate.  The  plug  at  the  side  k,  Fig.  60,  is  now  gradually 
withdrawn  and  the  accumulated  slime  is  carried  down  to  the  acid 
tank  o.  The  plug  is  replaced  and  the  compartment  is  then 
filled  with  water  in  which  the  zinc  is  again  rinsed  and 
rubbed,  and  the  loosened  precipitate  again  drawn  off.  About  three 
such  washes  are  enough  to  free  the  shavings  from  short  zinc  and 
from  precipitate.  Each  compartment  having  been  thus  cleaned 
out,  the  launder  is  cleaned  with  a  hose,  zinc  from  the  end  com- 
partments distributed  to  the  head,  and  the  rest  of  the  box  replen- 
ished with  fresh  shavings. 

When  all  the  boxes  have  been  cleaned  the  precipitate  is  al- 
lowed to  settle  for  a  short  time  in  the  acid  tank  and  the  clear 
liquid  is  then  siphoned  off  to  the  settling  tank,  see  Fig.  60.  The 
acid  tank  o  is  represented  in  detail. 

This  tank  is  provided  with  an  agitator,  which  can  be  kept  in 
motion  by  power  transmitted  from  the  line  shaft,  shown  in  the 
elevation  of  Fig.  59.  It  insures  thorough  agitation  of  the  sludge 
or  precipitate  when  undergoing  the  operation  next  to  be 
described. 

The  acid  treatment. — Upon  the  watery  slime  about  30  Ib.  of 
sulphuric  acid  is  poured  which,  acting  on  the  short  zinc,  pro- 
duces a  violent  effervescence.  As  this  subsides  the  slime  is 
stirred  by  the  agitator.  When  the  action  decreases,  about  15  Ib.  of 
hot  water  with  the  same  weight  of  acid  is  added  with  occasional 
stirring ;  and  so  on,  until  further  addition  of  acid  produces  but 


OF    THE    COMMON    METALS.  14! 

little  effervescence.  Then  the  mixture  is  allowed  to  stand  for 
two  hours,  and  tested  with  a  little  acid  to  see  that  decomposition 
is  complete.  The  total  time  for  this  operation  is  from  four  to  six 
hours. 

The  black-looking  mixture  is  now  diluted  with  hot  water  to 
within  a  few  inches  of  the  top  of  the  tank  and  pumped  through 
the  filter-press.  This  washing  with  hot  water  is  repeated  several 
times,  until  the  zinc  sulphate  has  been  quite  removed.  Finally 
the  tank  is  hosed  out,  the  washings  being  passed  also  to  the  filter- 


FIG.  61.     CENTRIFUGAL  PUMP. 

press.  In  filtering,  the  filter  cloths  are  covered  with  filter  paper, 
so  that  the  precipitate  does  not  touch  the  cloth,  this  paper  being 
afterward  burned,  and  the  ashes  mixed  with  the  precipitate. 
Fig.  62  represents  the  filter-press  used.  It  consists  of  a 
series  of  cast-iron  flat  frames  (each  frame  being  recessed),  be- 
tween which  the  filter  cloths  are  interposed.  The  liquid  enters 
the  press  under  pressure  from  a  pump  through  the  pipe  shown 
at  the  head  end  at  the  centre,  the  filtrate  escaping  by  bronze  cocks 
into  a  launder.  Leaning  against  the  launder  are  to  be  seen  two 
frames  showing  their  grooved  surfaces  by  which  the  filtrate  es- 
capes to  the  stop-cock.  To  the  left  is  shown  this  grooving  in 
detail,  the  dotted  line  on  the  section  showing  how  the  filter  cloth 


I  4-^  THE   MKTALLUKCY 

lies  against  the  frame.  The  precipitate  collects  in,  and  fills,  the 
recesses,  and  is  freed  from  much  of  the  remaining  moisture  by  air 
introduced  after  filtration.  The  frames  are  now  opened,  and  the 
precipitate  falls  or  is  scraped  from  the  frames  into  a  pan  set 
below. 

Drying  and  final  treatment  of  the  precipitate. — The  product  is 
transferred  to  a  cast-iron  pan,  44  by  24  in.  by  4  in.  deep.  This 
pan  is  placed  in  a  cast-iron  muffle-furnace,  where  it  is  gradually 
dried,  and  finally  heated  to  a  just- visible  red.  The  pan  is  then 
removed,  allowed  to  cool,  and  its  weighed  contents  mixed  with 
half  its  weight  of  a  flux  composed  of  borax  4  parts,  soda  2  parts. 
and  sand  I  part.  These  proportions  will  vary  slightly  at  different 
plants.  From  the  pan  the  material  is  cautiously  put  into  large 
plumbago  crucibles,  which  are  set  in  a  wind  furnace,  and  packed 
round  with  coke,  the  melting  being  done  as  in  silver  milling 
(Section  40,  which  see).  The  molten  metal  is  poured  into  conical 
molds,  and,  on  cooling,  the  slag  is  removed  and  the  gold  re- 
melted  into  an  ingot. 

The  Tavern cr  process. — In  this  process,  the  acid  treatment 
which  is  said  to  occasion  some  loss  of  gold,  is  done  away  with, 
and  the  gold  is  obtained  by  the  Taverner  method  as  follows : 
The  clean-up  of  the  gold-slime  is  conducted  as  described  above, 
the  precipitate  being  collected  in  the  acid  tank,  now 
called  the  clean-up  tank.  The  precipitate  is  pumped  to 
the  filter-press,  the  short  zinc  which  remains  at  the 
bottom  of  the  tank  being  heaped  to  one  side  and 
allowed  to  drain.  The  filter-press  slime  is  dried  for  a  few 
minutes  in  an  oven.  Thus  dried,  it  is  passed  through  a  4-mesh 
sieve,  weighed  and  mixed  with  fluxes  in  the  following  propor- 
tions. Slime  100  parts,  slag  20  to  30  parts,  sand  5  to  10  parts, 
litharge  60  parts.  This  is  placed  on  the  hearth  of  a  cold  rever- 
beratory  furnace  ( see  Fig.  63  I . 

Upon  this  is  placed  the  short  zinc  from  the  clean-up  tank,  then 
a  thin  layer  of  150  parts  of  litharge,  and  finally  a  thin  layer  of 
20  parts  of  slag.  A  slow  fire  is  started,  and  after  two  hours,  in- 
creased to  melt  the  charge,  which  takes  about  four  hours  longer. 
It  is  then  well  stirred,  mixing  in  some  sawdust  to  reduce  the  lith- 
arge; and  the  stirring  continued  until  the  slag  shows  clear  on  the 


OF   THE    COMMON    METALS. 


143 


144  THE     METALLURGY 

rabble.  The  slag  is  skimmed  off,  and,  after  standing  awhile  in  the 
pots,  is  poured  out,  the  shells  being  removed  for  re-melting.  The 
lead  surface  thus  being  cleaned,  the  zinc  burns  off  and  the  rich 
lead,  clean  and  soft,  is  ladled  into  bars.  These  bars  are  then 
treated  in  an  English  cupelling  furnace,  using  a  test  made  of 
bone  ash  (see  Fig.  152),  by  which  impurities  such  as  copper  are 
removed ;  and  finally  the  bars  are  fined  quite  as  described  in  Sec- 
tion 109.  The  litharge  produced  is  used  for  the  next  charge  to  be 
treated. 

The  alleged  advantages  of  the  process  are  ( I )  saving  in  the 
cost  of  treatment  (5.5  c.  against  24  c.  per  oz.  of  fine  gold)  ;  (2) 
absence  of  by-products;  (3)  increased  recovery  of  the  gold  (10% 
more  than  by  the  acid  process)  ;  (4)  facility  of  treating  foul  slag. 

28.     SECOND  METHOD  OF  CYANIDING.     SOUTH  AFRICA. 

South  African  practice,  brought  today  to  a  great  perfection, 
has  been  a  development.  At  first,  ores  containing  $9  values, 
were  milled  by  wet-stamping  and  amalgamating,  the  tailing,  still 
containing  some  $3.50  in  gold  value,  having  been  run  to  waste. 
With  the  introduction  of  the  cyanide  process  the  tailing  has  been 
cyanided,  and  further  values  cheaply  extracted.  Since  the  ore 
had  previously  been  stamped,  a  considerable  quantity  of  slime 
was  necessarily  made,  which  would,  if  allowed  to  remain  in  the 
ore,  practically  prevent  leaching.  As  a  first  step  in  cyaniding, 
the  slime  had  accordingly  to  be  removed  by  means  of  separators 
and  run  to  waste.  The  remaining  sand  was  then  treated  as  in 
the  first  method  already  described.  Later,  however  a  method  was 
devised  for  treating  slime. 

The  ore  called  'banket'  is  a  conglomerate,  the  gold  occurring 
in  the  cementing  material  which  unites  the  pebbles,  and  therefore 
fine  crushing  is  not  needed. 

After  coarse  crushing  the  ore  through  Blake  or  Gates  rock- 
breakers  it  goes  to  the  stamps  which  weigh  1,050  to  1,200  lb.,  and 
which  yield  4  to  6  tons  ore  per  stamp  crushed  through  a  3O-mesh 
screen,  with  the  use  of  about  8  tons  of  water  per  ton  of  ore. 
The  pulp  from  the  mill  flows  over  amalgamated  apron-plates,  as 
in  ordinary  gold  milling,  and  thence  to  a  tailing-wheel  and  to 


OF    THE    COMMON    METALS. 


145 


r 


I 


146  THE    METALLURGY 

a  group  of  pointed  box  classifiers  .'located  above  the  leaching 
tanks.  Some  of  these  wheels  are  as  large  as  50  ft.  in  diameter, 
they  are  very  efficient  elevators,  cheap  to  operate  and  are  made 
necessary  by  the  flat  ground  on  which  most  of  the  plants  are 
built. 

The  classified  coarse  sand  and  sulphide  pass  to  collecting  tanks, 
in  two  tiers,  one  superimposed  above  the  other.  They  are  made 
of  steel,  and  are  30  ft.  diam.  The  upper  or  collecting  tank  is 
usually  made  a  little  shallower  than  the  lower.  It  is  supported 
on  steel  I-beams  which  are  in  turn  supported  on  steel  or  cast- 
iron  columns  extending  through  the  lower  tank  and  encased  in 
a  protecting  cylinder.  The  sand  and  a  remnant  of  the  slime 
were  formerly  fed  into  the  upper  tank  by  means  of  a  pipe  dis- 
tributor, but  this  is  now  done  with  a  large  hose  which  is  moved 
about  at  intervals  by  hand.  The  upper  tank  is  provided  with 
circular  discharge  gates,  through  which  it  is  emptied  to  the 
lower  tank.  In  some  instances  the  upper  tank  is  merely  used 
for  collecting  the  sand;  in  others,  the  charge  is  given  a  very 
dilute  solution  of  cyanide  and  the  leaching  is  allowed  to  pro- 
ceed until  the  moisture  is  displaced.  In  the  lower  tank  the  sand 
is  leached  with  a  stronger  solution,  and  precipitation  carried  on 
by  means  of  zinc  shavings  or  electrolytically.  In  some  plants 
where  a  perfect  separation  between  sand  and  slime  is  effected 
the  whole  operation  of  settling  and  leaching  is  carried  on  in  the 
same  tank.  The  best  example  of  this  is  seen  at  the  cyanide  works 
of  the  Homestake  company. 

The  third  product  or  slime  is  now,  as  already  explained,  treated 
by  decantation.  The  slime  is  settled  in  large  tanks,  the  water 
drawn  off,  very  dilute  cyanide  solution  added,  and  the  agitation 
carried  on  by  transferring  the  pulp  from  one  tank  to  another  by 
centrifugal  pumps. 

South  African  practice  has  recently  felt  the  influence  of  West 
Australian  methods,  and  is  slowly  adopting  fine  grinding  of  the 
coarser  sand. 

Filter-pressing  is  practised  with  success  in  Australia  in  the 
treatment  of  roasted  ore.  The  Dehne  press,  resembling  Fig.  62, 
is  most  commonly  used.  These  are  arranged  in  series, 
and  receive  the  slime  and  fine  sand  after  treatment  by 


OF   THE    COMMON    METALS. 


147 


148 


THE    METALLURGY 


cyanide  solution.  It  was  at  first  customary  to  treat 
the  product  in  the  press,  and  to  full  the  latter  by 
means  of  compressed  air.  The  dissolving  of  values  is  now  done 
in  agitating  vats  and  the  press  rilled  direct  by  plunger  pumps — so 
that  now  the  press  is  merely  used  for  separating  the  solution 
from  the  slime  and  for  washing  purposes. 

In  the  treatment  of  slime  produced  in  wet-crushing,  we  have 
to  deal  with  a  less  permeable  material.  This  slime  is  treated 
by  presses  in  New  Zealand;  and  at  the  Homestake  mine.  South 
Dakota,  very  large  presses  with  a  capacity  of  25  tons  each  are 
being  used  to  treat  about  1,200  tons  of  slime  per  day.  Extraction 
is  carried  on  in  the  press  itself,  which  is  finally  emptied  by  means 
of  water  under  high  pressure  introduced  through  a  special  de- 
vice into  each  chamber  of  the  press.  This  does  away  with  one 
important  item  in  the  cost  of  operating,  namely  the  labor  of 
opening  and  closing  the  press. 


29.     THIRD  METHOD.     MEXICO. 

At  El  Oro,  Mexico,  the  ore  is  crushed  in  stamps,  and  as  much 
as  can  be  economically  .reduced  to  a  slime  is  treated  by  decan- 
tation.  The  fine  crushing  is  done  in  tube-mills  (Fig.  64),  and  the 
very  fine  sand  leached. 

At  the  Liberty  Bell  mill,  Tellu- 
ride,  Colo.,  the  product  of  80 
stamps  is  crushed  in  a  cyanide  so- 
lution, amalgamated,  and  then 
re-ground  in  tube-mills.  The  whole 
tube-mill  product  is  then  treated  in 
Hendryx  agitators  (Fig.  65),  after 
which  it  passes  to  the  Moore  filter- 
ing plant  for  filtration.  The  solu- 
tions are  precipitated  on  zinc  shav- 
ings. 

Fig.  66  illustrates  a  slime-plant 
for  the  treatment  of  a  wet-stamped 
ore  which  flows  into  the  settling- 
vat  A,  as  fast  as  it  is  produced.  FIG.  65.  HENDRYX  AGITATOR. 


OF   THE    COMMON    METALS. 


149 


I5O  THE    METALLURGY 

When  the  vat  is  full,  the  flow  goes  to  another  similar  vat  while  the 
contents  of  the  first  vat  are  allowed  to  settle,  milk-of-lime  being 
added  to  promote  settling.  The  supernatant  water  is  decanted  as 
fast  as  it  clears  while  the  thickened  slime  is  withdrawn  at  the 
bottom  into  the  agitating  vat  B.  Cyanide  solution  of  0.3%  is  now 
run  in  from  the  stock  vat  just  beyond  the  filter-press  C,  and  the 
whole  is  circulated  for  a  period  of  10  hours  by  aid  of  a  centrifugal 
pump  D,  which  withdraws  it  from  the  bottom  of  B  returning  it 
by  a  pipe  to  the  top  of  the  vat.  Besides  this  air  is  pumped  through 


FIG.  67.     PERSPECTIVE  VIEW  OF  WOODEN  LEACHING  VAT. 

the  solution  by  means  of  a  pipe  leading  to  the  apex  of  the  conical 
bottom.  Finally  the  pulp  is  pumped  to  the  filter-press  C,  the 
filtrate  going  to  the  gold  solution  vat  on  the  right.  From  this  it 
flows  through  the  zinc  boxes  E  E,  and  then  to  the  sump  situated 
at  the  lowest  level.  The  sump  solution  is  returned  to  the  stock  vat 
as  it  accumulates. 

The  Leaching  Vat. — This  may  be  made  either  of  wood  or  of 
steel.  In  hot  countries  the  steel  vat  is  to  be  preferred ;  but  in 
cold  countries  where  it  is  necessary  to  house  the  plant,  the  wood 
vat  gives  satisfaction.  The  latter  is  cheaper  in  first  cost  and 
easier  to  set  up ;  the  steel  vat,  on  the  other  hand,  is  less  liable  to 
leakage,  and  is  free  from  alleged  losses  due  to  absorption.  On 
the  other  hand,  the  steel  vat  costs  more  to  maintain,  as  it  requires 
a  periodical  coat  of  protective  paint  to  preserve  the  steel  from 
the  action  of  the  solution.  In  western  America  the  use  of  wooden 


OF    THE    COMMON    METALS. 


152 


THE    METALLURGY 


vats  predominates,  although  steel  vats  are  used;  the  latter  are 
favored  in  South  Africa  and  West  Australia.  Each  possesses 
its  advocates ;  but  in  general,  the  steel  vat  is  to  be  preferred. 


FIG.  69.     FILTER  BOTTOM   OF  WOODEN  LEACHING  VAT. 

Fig.  67  is  a  perspective  view  of  a  wooden  vat,  where  will  be 
seen  the  hinged  bottom  discharge  opening  through  which  the 
tailing  may  be  shoveled  or  sluiced. 

Fig.  69  gives  in  plan  and  in  section  the  construction  of  one  form 
of  filter-bottom.  It  consists  of  a  wooden  ring  2  in.  high  by  2.5 


OF   THE    COMMON    METALS.  153 

in.  wide  nailed  to  the  bottom  of  the  tank  I  in.  from  th£  side. 
Cleats  i  by  2  in.  are  nailed  flat  upon  the  bottom  of  the  tank  one 
foot  apart.  On  these  are  nailed  I  by  4  in.  strips  one  inch  apart. 
Upon  this  false  bottom  is  laid  cocoa  matting,  and  over  this  8-oz. 
canvas  filter-cloth  cut  12  in.  larger  in  diameter  than  the  inside 
of  the  tank,  the  edges  of  the  cloth  being  held  down  by  a  rope 


(b) 
FIG.  70.     PLAN  OF  STEEL  LEACHING  VAT. 

driven  with  it  into  the  i-in.  space  between  the  wooden  ring  and 
the  staves.  Sometimes  this  ring  is  omitted,  the  joint  being  made 
by  nailing  a  strip  around  the  tank  over  the  edge  of  the  canvas, 
as  shown  at  the  section  at  the  right  of  Fig.  69.  The  manner  of 
securing  the  edge  of  the  cloth  is  shown  more  in  detail  in  Fig.  70. 
Fig.  70  and  71  represent,  in  plan  and  in  section  respectively, 
the  construction  of  a  steel  vat,  having  a  perforated  board  bottom, 
convenient  in  case  one  intends  to  remove  the  tank  for  use  in  an- 
other place.  A  ring  of  flat  iron,  0.5  by  2.5  in.,  is  riveted  to  the 
side  of  the  tank,  leaving  a  space  of  0.75  in.  The  cleats  upon  the 
bottom  are  2  by  1.5  in.  laid  on  edge,  while  the  perforated  boards 
are  bored  with  0.75  in.  holes,  and  are  screwed  to  the  cleats.  As 
in  the  case  of  wooden  vats,  cocoa  matting  is  laid  upon  this  false 


154 


THE    METALLURGY 


bottom,  and  that  again  covered  with  a  filter  cloth  of  8  to  10  oz. 
canvas  duck,  whose  edges  are  calked  with  a  rope  into  the  0.75 
in.  space  between  the  sides  and  the  ring,  as  shown  at  g  in  the  sec- 
tional view.  Leaching  vats  are  made  of  varying  dimensions,  from 
1 6  to  50  feet  in  diameter  and  4  to  9  ft.  deep,  the  shallower  ones 
for  the  more  slimy  ores.  The  tank  is  usually  made  large  enough 
to  hold  one  day's  supply  of  ore,  to  ensure  uniformity  of  work  in 
thj  mill.  The  rate  of  percolation  is  also  very  much  increased 
by  the  use  of  the  vacuum  pump.  The  tailing  is  commonly  re- 
moved from  the  tanks  either  by  shoveling  them  out  by  a  side  or 


FIG.  71.     SECTION  OF  STEEL  LEACHING  VAT. 

by  a  bottom  opening ;  or  where  water  is  abundant  by  hosing  the 
residues  into  a  sluice  or  launder  marked  5"  in  Fig.  59  by  which 
they  are  carried  to  the  dump. 

Fig.  72  represents  a  side-  or  bottom-discharge  door  which 
is  bolted  to  the  outside  of  the  tank.  The  joint  of  the  door  is 
made  tight  with  a  rubber  gasket.  Fig.  73  represents  a  central 
bottom-discharge  valve  ( Fig.  73 )  operated  from  the  charging- 
platform,  very  convenient  when  the  ore  is  to  be  sluiced  out.  It  is 
bolted  securely  to  the  bottom  and  is  self-sustaining.  The  opening 
is  10  in.  in  diameter. 

The  zinc-bo.\~es. — Fig.  74  gives  a  plan,  an  elevation,  and  a  cross- 
section  of  a  set  of  wooden  zinc-boxes,  and  Fig.  75  is  a  per- 
spective view  of  three  sets  of  them  arranged  as  at  K,  K,  K,  Fig. 
60.  As  there  shown,  each  set  of  boxes  contains  nine  compart- 
ments, 12  by  15  by  24  in.  in  size.  These  compartments  have  a  per- 


OF    THE    COMMON    METALS. 


155 


forated  bottom  of  sheet-iron  or  of  wire-cloth  sustaining  the  zinc 
shavings  with  which  they  are  filled.  The  partitions  are  set  al- 
ternately down  and  up  to  compel  an  upward  flow  of  the  gold- 
bearing  solution,  so  as  to  bring  it  intimately  in  contact  with 
the  zinc  upon  which  it  is  to  be  precipitated.  In  the  elevation  a, 
the  entrance  of  the  solution  from  the  gold-solution  tanks  into  the 
first  box  and  its  discharge  to  the  sump  are  clearly  shown.  The 
boxes  are  set  at  a  grade  of  0.25  in.  per  foot ;  h  is  a  launder  into 
which  the  gold  precipitate  formed  in  the  boxes  can  be  at  the  time 


FIG.  72.     SIDE  DISCHARGE  DOOR. 

of  clean-up,  conveyed  to  the  launder  m  Fig.  60,  and  thence  to  the 
acid  tank  O.  Both  tank  and  launder  are  covered  with  lids  for 
the  protection  of  the  precious-metal  contents. 

For  small  plants  sheet-iron  individual  zinc-boxes  painted  with 
asphalt  paint  are  much  used  (see  Fig.  79).  In  cleaning  up  the 
boxes  can  be  removed  one  at  a  time  and  a  box  freshly  packed  with 
zinc  substituted. 

Fig.  61  represents  the  centrifugal  pump  used  to  return  the 
barren  solutions  from  the  sumps  to  their  respective  stock-tanks. 
It  is  operated  by  a  belt  from  a  counter-shaft,  shown  in  the. ele- 
vation at  Fig.  59. 

Roasting  as  a  Preliminary  to  Cvaniding. — We  have  already 
referred  to  the  fact  that  gold  may  occur  in  pyrite  as  a  film  on  the 
faces  of  the  crystalline  particles.  Such  ore  is  susceptible  to  direct 
cyaniding  when  the  ore  is  so  crushed  that  these  films  are  exposed 


156 


THE    METALLURGY 


to  the  action  of  the  solution.  If,  however,  such  ores  are  not  cya- 
nided  when  freshly  mined,  but  are  allowed  to  lie,  they  are  ren- 
dered acid  by  oxidizing  atmospheric  influences,  and  need  careful 
neutralization  by  means  of  caustic  soda  or  lime.  If  the  gold 
occurs,  finely  permeating  the  pyrite  crystals,  they  must  be  minutely 
ground  before  it  can  be  entirely  attacked  by  the  cyanide,  and  may 
then  be  treated  by  agitation.  Again,  where  the  gold  is  combined 
with  tellurium  to  form  calaverite  or  sylvanite, 
it  is  insoluble  in  cyanide  unless  roasted.  Roasting, 
moreover,  renders  an  ore  porous  in  texture  so 
that  it  is  more  easily  permeated  by  a  solution. 
We  thus  see  that  sulpho-tellurides  and  pyrite  ores 
in  which  the  gold  exists  in  the  substance  of  the 
crystals  are  both  eminently  suited  to  roasting  as 
a  preliminary  to  cyaniding.  The  extraction  and 
rapidity  of  leaching  is  so  improved  that  for  many 
ores  the  expense  of  roasting  is  quite  justified. 

30.     FOURTH  METHOD,  SOUTH  DAKOTA. 

The  ores  of  this  country  may  be  divided  into 
the  unoxidized,  or  blue  ores,  which  yield  but  little 
to  amalgamation  (20  to  30^  ),  and  into  the  dense 
oxidized  or  red  surface  ores,  from  which  some 
8or/r  may  thus  be  extracted.  The  gold  occurs 
evenly  distributed  and  in  a  finely  divided  condi- 
tion, such  that  it  may  be  recovered  by  cyaniding. 

The  ore  is  coarsely  crushed  by  rock-breaker  to 
1.5  in.  size  and  automatically  fed  to  a  battery  of 
FIG.  73.  looo-lb.  stamps  where  it  is  wet-crushed  to  20- 


the  solution  is  added  I  to  1.5  Ib.  sodium  hydrate 
to  neutralize  the  acidity  of  the  ore.  In  crushing  4  to  6  Ib.  quick- 
lime is  added  at  the  battery  per  ton  of  ore  to  assist  in  the  pre- 
cipitation of  the  slime. 

The  battery  discharge  is  elevated  by  sand  pumps  to  a  system 
of  cone  classifiers  (see  Fig.  76).  Here  it  is  delivered  to  a  distrib- 
uting-box 6  by  3  by  3  ft.  in  size,  which  serves  to  give  a 


OF    THE    COMMON    METALS. 


157 


C/J 


158 


THE    METALLURGY 


regular  supply.  The  flow  from  this  box  is  divided  evenly 
between  two  cone  settlers  which,  however,  have  no  upward  flow 
of  water,  so  that,  while  the  sands  settle  out,  they  still  contain  some 
slime.  The  spigot-discharge  of  the  two  cones  is  received  into  a 
distribution  launder  which  delivers  it  to  a  hydraulic  cone  classi- 
fier, having  the  usual  supply  not  of  water,  but  of  battery  solution 
already  mentioned.  Thus  the  sands  are  separated,  containing 


no  more  than  I  to  5%  of  slime,  while  the  slime,  amounting  to 
30%,  and  much  of  the  solution,  escape  together  over  the  top. 

Treatment  of  the  sand. — The  sand-tanks  are  filled  by  means  of 
the  Butters  &  Mein  distributor.  In  this  case,  however,  the  dis- 
tributor can  be  moved  by  an  overhead  trolley  from  tank  to  tank  as 
required  (see  Fig.  77).  The  sands  enter  the  empty  tank,  evenly 
filling  it.  the  accompanying  solution  draining  away  through  over- 
flow openings  and  through  the  filter-bottom,  so  that  there  is  no 
overflow  over  the  top  edge  of  the  tank.  It  takes  two  days  or 
more  to  fill  a  tank.  Additional  battery  solution  is  now  run  in, 
and  this  is  followed  by  barren  solution,  being  that  which  has  been 
passed  through  the  zinc-boxes.  This  is  followed  by  a  small 
amount  of  wash-water  of  o.i  to  0.2  ton  per  ton  of  sand.  The 
leaching  is  continuous  and  it  takes  at  least  8  days  to  leach  a 
single  charge.  There  is  no  strong  solution  properly  so  called. 
The  extraction  on  the  sands  amounts  to  about  75%. 

Treatment  of  the  slimes. — The  overflow  from  the  cones  goes  by 
launder  to  one  of  the  slime-tanks,  14  ft.  diam.  by  10  ft.  deep, 


OF    THE    COMMON    METALS. 


159 


shown  in  plan  an  elevation  in  Fig.  78,  and  goes  down  behind 
a  wooden  partition  extending  half  way  to  the  bottom  of  the  tank. 
Before  charging,  the  tank  is  filled  with  barren  solution,  the  slime 


FIG.  76.     CONE  CLASSIFIERS 


FIG.  77.     BUTTERS  &  MEIN  DISTRIBUTOR. 

is  run  in,  the  solids  settling  in  the  tank,  while  the  overflow 
goes  away  by  the  spout  e  and  any  foam  is  held  back  by  a  strip  of 
wood  s.  It  will  be  noticed  that  the  circular  launder,  surrounding 
the  tank,  is  not  now  used.  The  quick  settling-out  of  slime  is  due 
to  the  addition  of  the  lime  (4  to  6  Ib.  per  ton  of  ore)  at  the  bat- 


l6o  THE    METALLURGY 

tery.  The  overflowing  solution  is  strengthened  by  the  addition  of 
some  cyanide  and  sent  on,  as  battery  solution,  to  the  sand  treat- 
ment. The  flow  of  entering  slime  is  kept  up  until  the  solution, 
escaping  at  the  spout  e,  begins  to  look  turbid,  by  which  time  the 
tank  has  been  rilled  with  the  settled  material  to  the  depth  of  4  feet, 
when,  of  course,  the  entering  flow  is  transferred  to  another  tank, 
and  the  contents  of  the  just-filled  tank  allowed  to  settle  for  about 
10  hours.  As  the  slime  settles  the  supernatant  clear  liquid  is 
drawn  off  by  the  decanting  device  c.  This  is  gradually  lowered 
in  the  tank  until  finally  the  solution  is  decanted  to  within  an  inch 
of  the  settled  slime.  The  decanting  device  is  then  raised  and  a 
wash  of  barren  solution  run  in,  while,  at  the  same  time,  com- 
pressed air  is  driven  through  the  solution,  rising  through  it  by 
the  perforated  pipe  d.  This  is  allowed  to  settle  and  the  solution 
decanted  as  before,  the  operation  being  repeated  4  to  6  times  with 
barren  solution,  and  once  with  water.  Each  wash  consists  of  40 
tons.  After  this  the  slime,  containing  50%  moisture,  is  dis- 
charged through  a  bottom  gate,  indicated  at  h.  The  course  of  the 
solution  is  shown  in  the  accompanying  diagram.  It  will  be  seen 

Battery  Solution  Storage 


Batteries 


Cones 

7~\ 

Slimes  Sands 

r  ~r 

Slimes  Vats  Sands  Vats 

7      \  n      r 


Solution     Wash  water  goes 

Decanted  solution      Overflow  to  pre-    to  battery  solu- 

returned  to  bat-       while  filling          cipitating    tion  storage 
tery  solution  stor-    standardized          barrels, 
age.  and  then  to 

sand  vats. 

Barren  Solution 

goes  to  slime  vats,  sand  vats  and 
battery  solution  storage. 


OF    THE    COMMON    METALS. 


161 


Plan 


SCALE 


FIG.  78.     SLIMES  TANK. 


l62  THE    METALLURGY 

from  this  diagram,  that  there  is  no  escape  of  any  solution,  except 
that  which  gets  away  in  the  sand  and  slime  as  they  are  dis- 
charged from  the  tanks.  Xone  of  the  solutions  from  the  slime 
treatment  go  to  the  zinc-boxes  direct,  being  too  low  in  gold,  but 
pick  up  an  additional  load  of  the  latter  when  they  enter  the  sand- 
tanks,  either  directly  or  by  way  of  the  battery  storage  tank. 

Precipitation  of  the  gold. — In  place  of  the  ordinary  zinc-boxes, 
individual  barrels  or  tanks  may  be  used,  of  which  an 
illustration  is  given. 

The  gold  precipitate  is  subjected  to  the  acid  treatment,  already 
described  under  the  first  method,  and  yields,  in  South  Dakota 
practice,  a  bullion  of  450  to  600  fine.  The  total  extraction,  both 
from  sand  and  slime,  will  vary  from  68  to  75%,  the  potassium 
cyanide  consumed  will  amount  to  1.33  Ib.  and  the  zinc  used  up 
to  0.58  Ib.  per  ton  of  ore  treated. 

The  following  is  the  cost  of  treatment  per  ton  of  ore : 

Labor  $0.452 

Superintendence  0.090 

Cyanide  0.211 

Zinc  0.033 

Lime  0.012 

Power  0.226 

Shoes,  dies,  etc.  0.095 

Repairs  0.027 

Refining  0.030 

Assay  office  0.040 

General  expense  0.050 


$1.266 

31.     FIFTH  METHOD,  WEST  AUSTRALIA. 

Fig.  66  is  a  cross-section  of  a  complete  process  of  ore 
treatment  where  the  ore  is  re-ground  and  filter-pressed.  The  ore, 
while  it  may  contain  some  coarse  gold,  has  the  larger  part  of  it 
finely  divided,  so  that,  for  high  extraction,  it  needs  extremely  fine 
grinding  before  the  microscopic  particles  are  accessible  to  the 
cyanide  solution.  The  ore,  from  the  storage  bin  at  the  left,  is 


OF    THE    COMMON    METALS. 


automatically  fed  to  the  stamp  battery,  where  it  is  crushed  in  a 
cyanide  solution  of  0.1%.  The  coarse  gold,  if  any,  is  taken  out 
on  the  amalgamated  apron-plates,  whilst  the  pulp  is  raised  by  the 
elevator  to  two  classifying-cones,  which  remove  the  sand,  in  two 
grades,  the  coarse  going  to  one  table,  the  fine  to  another.  Any 
concentrate  is  removed  and  may  be  considered  to  have  sufficient 
value  for  shipping  away  to  the  smelter.  The  tailing  joins  the 
overflow  from  the  cones,  and  passes  to  the  slime  settler,  a  large 
cone  classifier,  which  removes  the  water  and  returns  it  to  the 


}    REDWOOD  MANUFACTURERS  COMPANY 


FIG.  79.     INDIVIDUAL  ZINC  BOXES  FOR  GOLD  PRECIPITATION. 

stamps,  while  the  settled  sand  and  slime  are  reground,  as  shown  in 
the  figures,  in  amalgamating  pans  without  the  use  of  mercury.  (In 
place  of  the  pan  a  favorite  machine  for  regrinding  is  shown  at 
Fig.  61.  It  consists  of  a  steel  riveted  cylinder  lined  with  flint 
stone,  making  22  to  27  rev.  per  min.,  5  ft.  diam.  by  20  ft.  long 
and  nearly  one-half  filled  with  flint  pebbles  the  size  of  one's  fist. 
The  sands  are  fed  in  at  one  end  by  a  hopper  and  are  ground  by 
the  pebbles  rolling  on, them  as  the  barrel  revolves.  The  ground 
material  escapes  at  the  other  end.)  Referring  again  to  the  dia- 
gram we  find  the  ground  material  from  the  pan  entering  the  agi- 
tating vat,  where  it  receives  a  farther  charge  of  cyanide  solution 


164  THE    METALLURGY 

of  o.22f/r ,  is  agitated  for  several  hours,  and  is  finally  run  ofif  to 
a  'montejus'  or  pressure  tank  5  ft.  diam.  by  12  ft.  high.  A  pipe 
from  the  bottom  of  the  tank  leads  to  the  filter-press,  and,  through 
this  is  driven  the  finely  ground  pulp  by  the  aid  of  compressed  air 
under  a  pressure  of  40  to  80  Ib.  per  sq.  in.  The  filtrate  from  the 
press  passes  as  usual  to  the  zinc-boxes.  A  charge  of  5  tons  of 
dried  slime  can  be  treated  at  a  time,  taking  upward  of  two  hours. 
When  ready,  the  press  is  opened  and  its  contents  drop  into  the 
tailing-hoppers,  which  deliver  to  conveying  belts  going  to  the 
dump. 

32.     THE  BROMO-CVAXOGEX  PROCESS. 

This  method,  also  called  the  Diehl  process,  is  practised  upon 
the  sulpho-telluride  ores  of  Western  Australia.  The  ore  carries 
iron  and  calcium  carbonates,  which,  after  roasting,  have  cement- 
like  properties  which  dispose  it  to  cake  or  set  together  when  sub- 
jected to  the  action  of  cyanide  solution  in  the  tanks.  Since  the 
Diehl  process  does  not  involve  roasting  of  the  ore,  it  has  come  in 
as  a  successful  method  of  treatment. 

It  has  been  found  that  to  get  a  good  extraction  from  the  ores 
they  must  be  finely  ground,  or,  as  it  may  be  termed,  slimed.  At 
the  same  time  we  must  distinguish  such  a  product,  which 
contains  many  hard  crystalline  particles  of  quartz,  from  a  real 
slime  made  up  of  colloid  substances  such  as  clay,  and  in  which 
no  hard  substance  exists.  Moreover,  the  gold  exists  only  in  part 
in  the  metallic  state,  the  rest  being  present  as  telluride  and  in  the 
sulphides.  Owing  to  the  brittleness  of  the  telluride,  the  finer  the 
slime  the  richer  it  is  in  gold. 

The  ore,  containing  2  to  3  oz.  Au  per  ton,  is  therefore  crushed 
through  a  stamp  battery,  and  passes  on  to  tube-mills  (see  Fig. 
68 )  where  is  it  crushed  so  finely  that  it  will  nearly  all  pass  a  200- 
mesh  screen.  The  product  from  the  tube-mills,  containing  5%  of 
sandy  particles,  goes  to  a  system  of  classifiers,  where  these  are 
separated  and  returned  for  re-crushing.  The  overflow  of  the 
classifiers  goes  again  to  spitzkasten  from  the  bottom  of  which 
is  drawn  off  the  thickened  pulp  containing  40  to  50^  of  solid 
matter  which  is  discharged  to  the  agitator.  This  is  a  tank 
25  ft.  diam.  by  8  ft.  high,  capable  of  holding  125  tons  of 


OF   THE    COMMON    METALS.  165 

pulp,  and  in  which  the  pulp  is  agitated  by  stirrers.  When  the 
agitator  has  been  filled,  cyanide  of  potassium  is  added  until  the 
strength  of  the  solution  is  brought  up  to  0.22%  per  ton  of  ore, 
and  the  whole  is  stirred  for  1.5  hours.  This  is  followed  by  an 
addition  of  bromo-cyanogen  to  increase  its  strength  by  0.055% 
of  the  salt  per  ton  of  dry  slime.  The  stirring  is  kept  up  and  at 
22  hours  from  the  start  i  to  4  Ib.  of  dry  quicklime  per  pound  of 
dry  slime  is  added.  The  charge  is  now  run  to  a  montejus  and  then 


FIG.  80.     ZINC  LATHE. 

driven  to  the  filter-presses  by  compressed  air.  The  press-cakes 
are  washed  first  with  2  tons  of  a  weak  solution,  then  with  the 
same  weight  of  water.  They  are  then  blown  nearly  dry  by  send- 
ing compressed  air  at  80  Ib.  pressure  through  them  for  15 
minutes.  The  press,  being  opened,  the  cakes  are  discharged  into 
cars  standing  below.  The  clear  filtrate  goes  to  the  zinc-boxes ; 
the  weak  washes  return  to  the  weak  solution  tanks  to  be  used 
over  again.  As  compared  with  extraction  with  plain  cyanide 


1 66  THE    METALLURGY 

which  will  vary  from  41  to  62%,  there  is  an  increase  to  77  to  97% 
or  of  about  35%. 

On  account  of  the  high  value  of  the  ore,  filter-pressing  is  essen- 
tial. These  presses  have  a  capacity  of  5  tpns  each,  and  consist  of 
50  frames,  each  40  in.  square,  the  cake  of  slime  being  2.5  to  3  in. 
thick.  This  gives  a  total  area  of  22  sq.  ft.  per  frame,  or  a  total 
of  1,100  ft.  per  press. 

33.     PRACTICE  UPON  CRIPPLE  CREEK  ORES. 

The  phonolite  ores  treated  may  be  divided  into  two :  First,  the 
altered  surface  ores  with  no  tellurium  and  containing  free  gold, 
and,  second,  the  deeper-lying  ores  containing  the  gold  combined 
with  tellurium  as  sylvanite  and  calaverite,  together  with  some 
nearly  barren  pyrite.  The  surface  ores  are  treated  raw,  the 
telluride  ores  receive  an  oxidizing  roast  before  leaching. 

Either  kind  is  coarsely  crushed,  dried  and  then  finely  crushed, 
as  described  in  detail  in  the  chapter  on  crushing.  Surface 
ores  are  crushed  to  4O-mesh  while  the  tellurides  are  crushed 
to  3O-mesh,  since  they  are  to  be  subsequently  roasted  whereby 
the  ore  is  made  more  porous  and  accessible  to  the  solu- 
tion. The  roasting  of  the  pulverized  ore  is  done  in  one  of  the  me- 
chanical roasters,  the  Argall,  the  Ropp  straight-line,  the  Edwards 
or  the  Holthoff-Wethey.  The  latter  furnace  completes  the  cool- 
ing of  the  ore  upon  its  lower  hearth,  while  in  the  case  of  the 
other  furnaces,  a  cooling  apparatus  must  be  used  in  addition. 
This  cooling  apparatus,  in  some  cases,  consists  of  a  trough,  water- 
cooled  by  pipes  below,  in  which  the  ore,  mixed  with  10  Ib.  lime 
per  ton,  is  moved  along  by  means  of  flights  or  scrapers,  which  de- 
liver it  at  the  end  of  the  trough  to  a  storage  bin.  At  the  discharge 
end  the  ore  may  be  sprinkled,  just  before  it  leaves  the  trough,  with 
strong  cyanide  solution.  This  at  once  cools  it,  prevents  its  dust- 
ing, and  starts  dissolution. 

Connected  with  the  drying  and  the  roasting  furnaces  are  two 
large  fans  discharging  10,000  cu.  ft.  of  air  per  minute  to  an  ex- 
tensive system  of  flues  terminated  by  a  bag  house.  The  dust,  thus 
recovered,  is  of  higher  grade  than  the  original  ore.  It  may  be 
bricked  and  smelted. 


OF    THE    COMMON    METALS.  l6/ 

The  leaching  tanks  are  of  steel  50  ft.  diam.  by  6  ft.  deep,  having 
a  false  bottom,  constructed  much  as  is  shown  in  Fig.  70.  The 
bottom  is  made  of  I  in.  stuff,  bored  with  0.75  in.  holes,  set  radi- 
ally and  sloping  5°  to  the  centre  discharge  or  else  set  flat. 

The  bottom  is  covered  with  cocoa-matting,  and  this  again  with 
8  or  10  oz.  duck  or  cotton  canvas.  The  empty  tank  is  filled  6  in. 
above  the  false  bottom  with  strong  solution  of  0.7%.  Ore  filling 
from  2-wheeled  buggies  is  then  begun  and  solution  is  run  in  from 
below  so  that  ore  and  solution  rise  at  the  same  time  in  the  tank, 
the  whole  work  of  charging  taking  50  hours.  The  first  solution 
is  allowed  to  stand  on  the  ore  24  hours,  then  more  is  added  and 
drawn  off.  This  is  repeated  until  the  gold  in  the  ore  is  reduced 
to  33%  of  its  original  gold  contents,  taking  5  days.  The  weak 
solution,  of  25%  KCN,  is  then  used  in  the  same  way  for  3  days, 
after  which  the  charge  is  treated  to  several  water-washes,  drained 
and  sluiced  out  through  a  bottom-discharge  valve,  Fig.  73, 
or  is  shovelled  out  at  side-discharge  doors,  Fig.  72,  all 
this  needing  two  days  more.  The  total  time  is  conse- 
quently 12.5  days.  The  original  ore,  as  charged  to  the  vat, 
contained  0.90  oz.  Au  and  0.5  oz.  Ag  per  ton,  while  the  tail- 
ing retained  0.06  oz.  Au  and  0.2  oz.  Ag.  The  extraction  was 
accordingly  93.33%  of  the  gold  and  60%  of  the  silver,  while  the 
loss  of  cyanide  was  1.75  Ib.  cyanide  per  ton  of  ore  treated.  The 
screen  analysis  from  the  above  charge  showed 

Coarser  than  4O-mesh  19%  containing  0.07  Au 

40  to  200     "  40%  "         0.06  " 

100  to  200     "  16%  "         0.06  " 

Less  than  200     "  25%  "         0.07  " 

This  shows  that  extraction  was  very  even,  being  as  complete  on 
the  coarser  size  as  on  the  slime. 

Each  kind  of  solution  passes  to  its  own  gold  tank,  and  from 
thence  it  is  slowly  run  through  the  zinc-boxes  to  the  sumps,  the 
solution  being  then  strengthened  for  re-use.  The  zinc  consumed 
amounts  to  0.9  Ib.  per  ton  of  ore  treated. 

The  cost  of  treatment'  of  Cripple  Creek  ore  may  be  taken  as 
$2  to  $2.50  per  ton,  roasting  included,  when  performed  on  the 
large  scale  above  indicated. 


1 68  THE  METALLURGY 

34.     CYAXIDING  SULPHO-TELLURIDE  ORES. 

Cripple  Creek  ores  of  this  character  need  a  roast  before  they 
are  cyanided.  For  economic  working  a  mill  treating  such  ores 
should  have  a  capacity  of  at  least  400  tons  daily.  At  the  Golden 
Cycle  mill,  Colorado  City,  Colo.,  the  practice  is  as  follows: 

The  ore  is  coarsely  crushed  to  2-inch  cubes  in  rock-breakers, 
and  it  then  passes  through  heavy  crushing-rolls  wThich  reduce  it 
to  about  one  inch  size ;  from  the  rolls  the  ore  is  conveyed  to  auto- 
matic storage  bins,  with  capacity  of  8,000  tons.  The  ore  is  re- 
moved from  the  storage  bins  on  automatic  belt-conveyors,  which 
deliver  it  to  the  final  crushing  rolls,  where,  without  preliminary 
drying,  the  ore  is  crushed  to  quarter-inch  cubes  or  finer.  The 
coarsely  crushed  ore  is  next  delivered  to  the  roasters,  it  having 
been  found  that  on  ore  of  this  size  roasting  can  still  be  thoroughly 
done.  The  telluride  of  gold,  where  segregated,  roasts  into  shot 
or  particles  of  gold,  too  large  generally  for  solution  with  potas- 
sium cyanide,  hence,  the  roasted  ore,  after  cooling,  is  ground  in 
cyanide  solution  and  amalgamated  in  Chilean  mills  (see  Fig.  81). 

The  pulp  passes  over  amalgamated  plates  and  the  fine  grinding 
to  6o-mesh  is  completed  in  pans.  By  this  method  all  the  amal- 
gamable  gold  is  removed  and  the  sands  reduced  to  a  product 
of  low  tenor ;  the  pulp  is  next  lifted  to  classifying-cones,  where 
the  slime  is  separated  from  the  sand  in  the  usual  manner ;  the 
latter  being  conveyed  directly  to  the  leaching  vats,  while  the 
separated  slime,  after  four  hours'  agitation  in  a  conical-bottom 
tank,  is  drawn  off  to  filter-presses  for  final  treatment. 

35.     ACTIOX  OF  COPPER  ix  CYAXIDIXG. 

Copper  in  the  ore  is  generally  understood  to  be  a  serious  draw- 
back to  successful  cyaniding,  since  it  passes  into  the  solution, 
and  consumes  cyanide.  It  is  customary,  in  the  presence  of  a 
little  copper,  to  strengthen  the  solution  as  it  enters  the  zinc- 
boxes,  thus  preventing  its  precipitation  to  some  extent.  Copper 
always  goes  into  solution,  using  up  potassium  cyanide  in  so 
doing,  and  thus  increasing  the  cost  for  that  chemical. 

The  Hunt  ammonia-cyanide  process. — At  Dale,  San  Bernar- 
dino County,  Cal.,  a  copper-bearing  gold  ore  was  treated  as  fol- 


OF    THE    COMMON    METALS. 


i6g 


lows :  In  crushing,  8  Ib.  of  quick-lime  was  added  per  ton  of  ore, 
and  in  the  vats  the  ore  was  treated  with  a  0.15%  solution  of  potas- 
sium cyanide  to  which  had  been  added  6  Ib.  ammonium  chloride 
per  ton  of  solution.  This  was  allowed  to  remain  for  12  hours 
in  contact  with  the  ore,  after  which  the  latter  was  drained  and 
washed,  the  operation  requiring  6  days.  The  filtrate  was  run 


FIR.  81.     MONADNOCK  (CHILEAN)    MILL. 


through  the  zinc-boxes  as  usual.  When  using  the  ordinary 
cyanide  solution,  there  was  a  loss  of  8  Ib.  cyanide  per  ton  of 
ore  treated,  while  by  the  above  method,  this  was  reduced  to 
I  Ib.  per  ton.  The  method  has  also  been  successfully  applied  to 
the  treatment  of  copper-bearing  mill-tailing,  long  exposed  to  at- 
mospheric influences. 


I7O  THE  METALLURGY 

36.     PRECIPITATION  WITH  ZINC  DUST. 

Zinc  dust  is  a  product  of  zinc  smelting,  containing  90% 
metallic  zinc.  Its  use  as  a  precipitant  'has  been  conspicuously 
successful  at  the  Homestake  cyanide  mills  in  South  Dakota. 
The  gold-bearing  solution  is  alternately  treated  with  the  dust 
in  two  large  tanks.  When  a  tank  is  full,  an  emulsion  of  zinc 
dust  is  sprinkled  on  the  surface  of  the  solution  in  the  proportion 
of  about  l/4  Ib.  per  ton  of  solution.  This  is  forced  on  the  tank 
under  air  pressure  from  a  small,  conical-shaped  vessel,  in  which 
the  emulsion  is  prepared.  A  violent  jet  of  air  entering  through 
perforated  pipes  is  then  introduced  into  the  tank,  and  the  agita- 
tion kept  up  for  fifteen  minutes,  after  which  the  whole  content 
of  the  tank  is  elevated  by  a  large  pump  to  the  filter-press,  which 
it  enters  under  about  10  Ib.  pressure.  The  excess  of  zinc  dust  thus 
accumulates  in  the  press  until  the  clean-up. 


37.     EXTRACTION  OF  GOLD  BY  SMELTING. 


Gold  may  be  recovered  from  its  ore  by  the  processes  of  lead- 
silver  or  of  copper-matte  smelting,  practically  all  of  the  gold 
being  saved.  This  would  be  the  best  way  of  obtaining  gold 
from  its  ores  were  it  not  for  the  expense  of  smelting,  especially 
of  the  silicious  ores.  Thus,  the  charge  for  smelting  a  silicious 
ore  by  a  silver-lead  smelting  works  would  be  $10  per  ton,  while 
it  is  often  milled  for  less  than  $i  per  ton,  with,  however,  a  lower 
extraction  or  recovery  of  the  gold.  Then  again,  the  ore  if 
smelted  must  be  sent  to  the  nearest  smelting  works,  which  in- 
volves an  additional  expense  for  freight. 


PART  IV.    SILVER 


PART  IV.    SILVER. 

38.     SILVER  ORES. 

Silver  sometimes  occurs  native  in  flakes  or  plates,  as  wire- 
silver,  or  as  adherent  to  native  copper. 

The  principal  ores  of  silver  are  as  follows : 

Cerargarite  (horn  silver,  silver  chloride),  AgCl,  contains  when 
pure  75.3%  Ag.  It  is  widely  distributed  and  in  mines  is  found 
in  the  upper  oxidized  zones.  It  is  probable  that  much  of  the 
so-called  chloride  ore  is  in  reality  a  chloro-bromide  (embolite). 
The  ore  is  readily  amalgamated  or  free-milling. 

Argentite,  Ag2S,  contains  87.1%  Ag.  This  is  one  of  the 
most  common  of  silver  ores,  and  is  capable  of  being  acted  on  by 
chemicals.  In  this  respect  it  is  more  docile  than  the  sulphides 
associated  with  arsenic  and  antimony,  namely : 

Stephanite,  5  Ag2S,  Sb2  S3  containing  68.5%  Ag, 

Pyrargarite,  3  Ag2  S,  Sb2  S3  containing  59.8%  Ag.,  and 

Proustite  3  Ag2S  As2S3  containing  65.4%  Ag. 

Finally  we  have  silver  sulphides,  which,  besides  the  above- 
named  impurities,  have  copper  also.  They  are  polybasite  9  (Ag, 
Cu)  S  (SbAs)2  S3  with  64  to  72%  Ag,  and  fahlerz  (gray- 
copper  ore)  the  most  complicated  of  all,  4  Cu  Fe  Ag2  (Hg  Zn)  S, 
(SbAs)  S3,  in  which  the  silver  varies  from  0.06  up  to  31%, 
the  silver  existing  in  large  quantity  in  the  more  arsenical  and 
less  in  the  more  antimonial  varieties. 

Besides  the  above  widely  distributed  ores,  there  are  others 
less  common,  but  sometimes  important  in  certain  localities.  They 
are  dyscrasite  Ag3  Sb,  stromyerite  Ag2  S,  Cu2  S,  and  freiesle- 
benite  5(PbAg2)S. 

A  number  of  rare  minerals  containing  silver  could  also  be 
enumerated,  but,  for  the  purpose  of  the  metallurgist,  the  ones 
named  above  are  the  more  important  ones.  As  they  come  to 
him  they  contain  an  overwhelming  proportion  of  gangue,  so  that 


1/4  THE    METALLURGY 

many  of  them  are  treated  containing  no  more  than  o.i  to  0.2% 
of  silver.  Thus  we  may  have,  as  at  the  Comstock  mines,  the 
silver  in  native  form  and  as  sulphides  associated  with 
pyrite,  galena,  blende  and  chalcopyrite,  and  having  a  gangue 
of  quartz,  calcite,  dolomite  and  the  oxides  of  iron  and  man- 
ganese. At  the  Ontario  mine,  Park  City,  Utah,  the  silver  is 
found  in  argentite  and  gray-copper  ore,  associated  with  the 
heavy  minerals,  blende  and  galena,  and  having  a  gangue  of 
quartz  and  clay.  These  heavy  minerals  also  carry  some  of  the 
silver,  which  accompanies  them,  into  the  concentrate  wherever 
concentration  is  practised. 

39.     SILVER  MILLING  AND  AMALGAMATION. 

The  ores  of  silver  which  can  be  successfully  subjected  to  free 
milling  are  such  as  contain  the  silver  (and  gold)  in  a  form 
capable  of  being  acted  upon  by  mercury,  assisted  by  agitation, 
heat  and  certain  chemicals. 

The  operation,  after  stamping,  is  performed  in  amalgamating 
pans,  each  being  worked  for  several  hours,  and  the  reactions 
often  being  slow.  The  forms  of  silver  are  horn-silver  (chloride), 
free  silver  (native)  and  certain  silver  sulphides,  notably  argen- 
tite. Base  minerals,  such  as  pyrite,  calcopyrite,  galena,  blende, 
and  especially  arsenic  and  antimony,  interfere  with  the  success  of 
the  process  in  two  ways. 

1.  Directly  by  fouling  the  quicksilver,  and  also  by  checking 
the  reactions  of  the  chemicals. 

2.  By  carrying  off  in  combination  such  silver  as  they  contain 
not  capable  of  being  liberated  by  pan  amalgamation.     There  is 
no   sharp   line   between    free-milling   and    roasting-milling   ores. 
Often  the  upper  part  of  a  vein  is  free  milling,  and  the  lower  part 
contains  base   metals  and   sulphides,   so  that  finally   it  becomes 
necessary  to  roast  the  ore.     The  best  extraction  is  obtained  on 
the  decomposed  or  oxidized  ores,  from  which,  by  weathering,  the 
sulphur  has  been  removed,  and  where  the  silver  occurs  in  form 
capable  of  being  reacted  on  by  mercury.     Occasional  deposits  of 
silver  chlorides  and  of  native  silver,  together  with  the  oxidized 
ores,  make  up  all  the  ores  capable  of  free  milling. 


OF    THE    COMMON    METALS.  175 

Silver  milling  may  be  divided  into: 

1.  The  wet-silver  mill  or  Washoe  process,  including 

(a)  The  tank  mill, 

(b)  The  Boss  process. 

2.  The  dry-silver  mill,  or  Reese  River  process. 

3.  Combination  mills,  combining  the  gold  mill  and  the  wet- 
silver  mill. 

In  the  treatment  of  silver  ores  the  principal  thing  which  we 
note  is  the  time  required  to  effect  the  amalgamation.  Thus,  in 
gold  milling,  that  portion  of  the  gold  not  caught  on  the  inside 
plate  comes  in  contact  with  the  outside  amalgamated  apron-plate 
for  a  few  seconds  only.  In  silver  milling,  on  the  contrary,  the  pul- 
verized ore  is  ground  and  mixed  most  intimately  with  the  mer- 
cury for  hours  in  an  amalgamating  pan,  the  operation  being  as- 
sisted by  heat,  and  by  the  addition  of  chemicals  to  promote  the 
amalgamation. 

Again,  in  gold  milling,  an  ore  containing  0.5  oz.  Au  per  ton 
can  be  satisfactorily  milled.  In  silver  milling  an  ore  of  equivalent 
value  would  contain  20  oz.  Ag  per  ton  of  ore,  or  40  times  as 
much  metal  as  in  the  first  case.  Thus  it  can  be  seen  why  so 
much  time  and  pains  must  be  taken  in  silver  milling  to  be  sure 
that  all  the  metal  possible  shall  be  recovered.  Even  so,  there 
may  still  remain  several  ounces  of  silver  in  the  tailing. 

Chloridising  amalgamation  versus  free  milling. — The  amal- 
gamation of  roasted  silver  ores  gives  a  higher  extraction,  con- 
sumes less  iron  and  loses  less  quicksilver  than  is  the  case  in  wet 
milling.  On  the  other  hand  the  expenses  of  roast  amalgamation 
are  high,  and  the  volatilization  loss  is  large,  hence  the  Reese 
River  process  is  not  extending.  Probably  many  ores  may  give 
a  good  yield  after  a  quick  oxidizing  roast  in  place  of  a  slow 
chloridizing  roast. 

Roast  amalgamation  is  being  replaced  by  the  combination 
process,  in  conjunction  with  the  smelting  of  the  concentrate. 

The  lower-grade  silicious  silver  ores  are  now  being  treated 
with  marked  success  in  Mexico  and  elsewhere  by  a  combination 
of  concentration  and  cyaniding.  At  Tonopah,  Nevada,  the  cheap- 
est method  of  treating  the  silver-gold  ores  of  that  district  has 
been  found  to  be,  after  a  careful  investigation  of  other  methods, 


176  THE    METALLURGY 

wet  crushing  in  stamps,  close  concentration,  and  cyaniding  of 
sand  and  slime  after  the  usual  separation  in  classifiers. 

40.     THE  WET-SILVER  MILL  OR  WASHOE  PROCESS   (TANK 

SETTLING). 

The  process  is  suitable  to  so-called  free-milling  ores,  in  which 
the  silver  is  found  in  native  form  or  as  chloride.  Sometimes 
a  little  silver  sulphide  may  occur,  but  the  ore  should  be  free  from 
lead  and  from  any  tough,  clayey  gangue. 

Fig.  82  is  a  sectional  elevation  of  a  wet  crushing  silver  mill. 
The  ore  coming  from  the  mine  is,  as  in  gold  milling,  dumped  over 
a  grizzly  or  bar-screen  set  at  an  inclination.  As  a  result  the 
larger  pieces  of  ore  (oversize)  are  screened  out  and  fall  close 
to  the  mouth  of  the  crusher,  while  the  finer  portion  (undersize) 
passes  by  chute  direct  to  the  receiving  bin.  The  lump  ore  is  mean- 
while fed  to  the  Crusher,  and  joins  the  rest  in  the  receiving  bin. 
As  a  result  of  using  a  grizzly,  the  crusher,  relieved  from  the  fine 
ore,  can  put  through  more  material.  This  work  of  coarse  crush- 
ing should  all  be  done  during  the  ten  hours  of  the  day,  and  the 
receiving  bin  should  be  large  enough  for  holding  a  night's  sup- 
ply. Crushing  is  done  to  1.25  in.  size,  and  preferably  even  to 
0.75  in.,  since  by  finer  crushing  the  tonnage  of  the  stamps  may 
be  increased.  From  the  receiving  bin  the  ore  is  fed  to  the  stamps 
by  an  automatic  feeder.  The  mortar,  in  silver  milling,  has  a  double 
discharge  since  the  machine  is  intended  for  crushing,  and  the 
increased  screen-opening  gives  larger  tonnage.  The  screens  are 
of  3O-mesh.  (For  further  particulars  of  gravity  stamps  see 
the  detailed  description  under  gold  milling.)  Water,  (about 
8  tons  per  ton  of  ore)  is  fed  in  to  make  a  pulp,  which  flows  or 
is  splashed  through  the  screens  and  then  goes  by  launder  to 
the  settling  boxes.  The  settling  tanks  or  reservoir  consists  of  a 
space  divided  into  boxes  or  compartments,  each  9  ft.  long  5  ft. 
wide  and  3  ft.  deep.  The  flow  of  the  pulp- is  from  box  to  box 
until  it  finally  goes  by  launder  to  a  settling  pond  outside  the 
mill.  In  the  first  boxes  most  of  the  solid  matter  settles  out,  a 
further  portion  dropping  in  the  succeeding  boxes  until  the  still 
dirty  water  passes  to  the  settling  pond.  Here  it  has  its  final 


1/  THE    METALLURGY 

chance  to  settle  and  clear  itself,  after  which  the  settled  water 
runs  to  waste,  or  may  be  used  over  again  in  the  mill  if  the  \vater 
is  so  scarce  that  it  is  an  object  to  do  so,  while  the  pond  settlings 
are  treated  in  pans  as  they  accumulate.  When  the  first  settling 
box  is  full  it  is  by-passed  or  cut  out  by  turning  the  flow  into 
the  next  box.  The  full  box  is  then  shoveled  out  upon  the  floor 
adjoining,  and  taken,  as  needed,  to  be  fed  to  the  pans.  Mean- 
while, the  just  emptied  box  has  the  flow  of  the  last  one  turned 
into  it,  thus  forming  the  final  one  of  the  series.  The  launders 
are  so  arranged  that  all  this  can  be  conveniently  done. 

The  ore  is  now  ready  for  further  treatment  and  is  charged 
into  5-ft.  diam.  amalgamating  pans  together  with  mercury.  Fig. 
83  is  an  illustration  of  such  a  pan  as  is  used  today  in  silver  milling. 
It  consists  of  a  cylindrical  cast-iron  pan  furnished  with  a  central 
cone  through  which  a  shaft  rises  carrying  the  muller.  The 
muller  has  attached  to  its  lower  surface  6  cast-iron  shoes  2.5 
in.  thick.  The  muller  can  be  raised  or  lowered  as  desired  by 
means  of  a  hand  wheel  and  screw  on  top  of  the  shaft.  The 
bottom  of  the  pan  has  fixed  to  it  iron  plates  or  dies,  forming 
the  lower  or  fixed  grinding  surface.  The  muller,  which  revolves 
60  rev.  per  min.,  not  only  serves  to  grind  the  ore  more  finely 
but  also  to  divide  up  the  mercury  at  the  bottom  of  the  pan  and 
to  bring  the  small  globules  of  it  in  intimate  contact  with  the 
ore  pulp.  There  is  set  up  a  movement  or  current  of  the  pulp 
toward  the  periphery,  where  it  rises,  flows  toward  the  centre, 
and  sinks  down  near  the  centre  and  again  comes  under  the  muller 
to  be  again  ground  and  brought  in  contact  with  the  mercury. 
To  assist  this  action,  the  rising  pulp  is  deflected  by  means  of  cast- 
iron  wing-plates.  The  reaction  in  the  pan,  by  which  the  silver 
in  the  pulp  becomes  amalgamated,  proceeds  most  favorably  when 
the  pulp  is  heated  nearly  to  boiling  by  means  of  steam  generally 
introduced  through  a  pipe  which  dips  beneath  the  surface 
of  the  charge.  Sometimes  also  the  pan  is  provided  with  a  double 
bottom,  into  which  exhaust  steam  from  the  engine  may  be  intro- 
duced. 

The  reactions  which  take  place  in  the  pan  are  as  follows  : 
Native  silver  in  threads,  films  or  particles  is  readily  taken  up 
and  dissolved  in  mercury,  forming  an  amalgam  which,  mixed 


OF    THE    COMMON    METALS. 


179 


with  a  large  excess  of  mercury,  resembles  it.  Silver  chloride 
coming  in  contact  with  mercury  is  decomposed  by  the  latter  thus : 
2  AgCl  +  2Hg  =  Hg2Q2  +  2  Ag,  the  silver  then  amalgamat- 
ing with  other  mercury.  But  the  Hg2  C12  does  not  long  remain 
as  such  since  the  iron  bottom  of  the  pan  and  particles  of  iron 
abraided  from  the  stamps  soon  decompose  it  again,  settling  free 
the  mercury.  In  many  so-called  free-milling  ores  there  is,  how- 


FIG.  83.     COMBINATION  AMALGAMATING  PAN. 

ever,  some  silver  sulphide  which  is,  in  part,  decomposed  by  the 
mercury  as  follows: 

Ag2  S+2  Hg=Ag2  Hg+HgS,  the  latter  salt  Being  lost. 
Chemicals  are  also  added  to  promote  decomposition  of  the  silver 
sulphide,  namely:  copper  sulphate  and  common  salt.  From  6 
to  1 8  Ib.  of  salt  and  3  to  9  Ib.  of  copper  sulphate  are  added,  per 
ton  of  ore  treated.  The  reactions  which  take  place  are,  as  gen- 
erally given. 


l8o  THE    METALLURGY 

Cu  SO4-(-2  Na  Cl=Na2  SO4-|-Cu  C12 ;  and  this  latter  compound 
acting  on  the  silver  sulphide  decomposes  it:  Ag2  S+Cu  C12= 
CuS-f-2  Ag  Cl,  which  latter  is  then  decomposed  by  the  mercury 
as  already  stated. 

In  working  the  pans  the  mullers  are  first  raised  half  an  inch, 
water  is  run  in,  the  muller  set  in  motion  and  3,000  Ib.  ore  charged 
in.  The  ore  and  water  are  stirred  to  a  pulp  of  about  the  con- 
sistency of  honey,  and  nearly  fill  the  pan.  The  muller  is  now  low- 
ered until  the  shoes  touch  the  dies,  and  the  ore  is  finely  ground 
for  about  one  and  one-half  hours,  the  contents  of  the  pan  being 
meanwhile  heated  up.  The  shoes  are  then  raised,  300  Ib.  or  10% 
of  mercury  is  added,  and  the  mixing  is  continued  for  four  hours. 
Amalgamation  proceeds  rapidly  at  first,  but  soon  diminishes. 
Thus,  at  the  end  of  the  first  hour  74.7%  of  the  silver  was  amal- 
gamated, at  the  end  of  the  second  76.3%,  at  the  end  of  the 
third  77-7%  and  in  four  hours  81.0%,  after  which  no  more  was 
taken  up.  The  amalgamation  being  complete,  the  pan  is  emptied 
to  the  settler.  About  15  minutes  before  discharging,  the  speed 
of  the  muller  is  reduced  to  40  revolutions,  and  the  pan  filled 
to  the  top  with  water.  A  plug,  closing  the  discharge  opening 
at  the  side  of  the  pan,  is  now  pulled  out  and  the  entire  contents 
of  the  pan  run  to  the  8-ft.  settler,  the  amalgam  being  separated 
there  from  the  ore.  Emptying  takes  half  an  hour,  water  being 
allowed  to  run  into  the  settler  at  the  same  time. 

The  complete  separation  of  the  amalgam,  mixed  of  course 
with  a  large  amount  of  mercury,  is  effected  in  the  settler,  there 
being  one  settler  provided  for  each  two  amalgamating  pans. 

The  settler  is  a  cylindrical  vat  or  pan  8  ft.  diam.  by  3  ft.  deep 
with  a  cast-iron  bottom  and  wooden  sides.  As  shown  in  Fig.  84, 
it  resembles  an  amalgamating  pan,  having  a  muller  provided 
with  slioes  of  wood,  since  the  object  of  the  pan  is  to  stir,  not  to 
grind.  The  shoes  nearly  touch  the  bottom  of  the  settler,  and 
can  be  raised  or  lowered,  as  desired,  like  the  muller  of  the  amal- 
gamating pan.  The  grooved  border  of  the  bottom  of  the  settler 
slopes  to  the  side,  conducting  to  an  outlet  which  is  provided  with 
a  mercury-well.  The  mercury,  settling  out,  collects  in  this  well 
to  a  height  to  balance  the  contents  of  the  pan,  say  two  to  three 
inches.  At  different  heights  in  the  side  of  the  pan  openings  are 


OF   THE    COMMON    METALS. 


181 


provided,  closed  by  plugs,   for  the  discharge  of  the  amalgam- 
free  tailing. 

The  muller  of  the  settler,  being  in  motion  at  the  rate  of  15 
rev.  per  min.  with  the  shoes  8  in.  above  the  bottom,  the  content 
of  the  two  pans  is  run  into  it,  and  water  is  added  to  6  in.  of  the 


FIG.  84.     EIGHT-FOOT  SETTLER. 


top.  After  half  an  hour  the  shoes  are  gradually  lowered,  nearly 
touching  the  bottom  by  the  end  of  two  hours.  The  purpose  of 
the  agitation  is  to  keep  the  lighter  portion  of  the  ore  (now  called 
the  tailing)  in  suspension,  while  the  amalgam,  mercury,  sulphides 
and  fine  particles  of  iron  collect  at  the  bottom  of  the  settler. 
The  stirring  is  kept  up  for  3.5  hours,  after  which  the  highest 


1 82 


THE    METALLURGY 


plug  at  the  side  of  the  settler  is  removed,  and  the  turbid  water 
allowed  to  flow  away  by  launder.  The  succeeding  plugs  are  then 
one  by  one  withdrawn  until  the  settler  is  emptied  of  all  except 
the  heavier  portion  above  mentioned.  Emptying  lasts  half  an 
hour.  Since  the  escaping  tailing  contains  some  portion  of  sul- 
phides and  particles  of  amalgam,  it  is  often  run  over  concen- 
trating tables  to  recover  this  valuable  heavier  portion,  other- 
wise it  runs  to  waste.  The  amalgam  mixed  with  mercury,  or, 

as  it  may  be  called,  diluted  amalgam, 
collects  in  the  mercury-well,  overflow- 
ing by  a  higher  opening  to  the  amal- 
gam safe,  Fig.  85.  Here  it  enters  a 
conical  canvas  sack  or  filter,  the  mer- 
cury oozing  through  the  pores  of  the 
canvas,  while  the  amalgam,  containing 
say  14%  Ag,  is  retained.  Occasion- 
ally the  sack  is  squeezed  to  remove 
the  surplus  mercury,  and  the  amal- 
gam, containing  20  to  28%  Ag,  is 
reserved  for  retorting.  The  mercury 
is  drawn  off,  or  flows  away  by  the 
outlet  p  to  be  used  again.  In  the 
large  mills  the  mercury  is  returned  to 
a  cast-iron  tank  placed  at  a  level  above 
the  pans,  being  raised  by  a  mercury 
elevator  which  is  an  ordinary  belt  ele- 
vator furnished  with  iron  buckets. 

The  loss  of  mercury  is  from  I  to  3.5  Ib.  per  ton  of  ore,  and  is 
greater  in  talcose  or  clayey  ores,  and  in  those  carrying  cerrusite 
or  sulphides  of  copper  and  lead.  It  is  also  lost  by  grease,  coat- 
ing over  globules  of  mercury,  and  by  flouring  caused  by  the 
stamping,  grinding  or  washing. 

Treatment  of  the  amalgam. — Since  the  quantity  of  metal  re- 
covered in  silver  milling  is  so  much  greater  than  in  gold  milling, 
the  retorting  of  the  amalgam  must  be  performed  on  a  larger 
scale.  Fig.  86  shows  sections  of  a  combined  retorting  and  melt- 
ing furnace  for  a  silver  mill  or  for  a  large  gold  mill,  showing 
also  the  arrangement  for  handling  the  melting  crucibles.  The 


FIG.  85.     AMALGAM  SAFE. 


OF    THE    COMMON    METALS. 


retort,  as  shown,  consists  of  a  cast-iron  cylindrical  vessel  4  to  5 
ft.  long,  10  to  14  in.  diam.,  of  cast  iron  1.5  in.  thick,  and  weighing 
900  Ib.  It  is  provided  at  one  end  with  a  water-cooled  pipe, 


FIG.  86.    HORIZONTAL  RETORT  AND  MELTING  FURNACE  FOR  SILVER  MILL. 

the  end  dipping  in  the  water  of  the  vessel  provided  to  collect  the 
mercury.  The  other  end  is  provided  with  a  cover  or  retort-head, 
securely  and  tightly  clamped  by  a  screw. 


184  THE    METALLURGY 

Operation. — The  charge  of  amalgam,  containing  20%  of 
mercury,  should  not  more  than  half  fill  the  retort,  and  weighs 
1,000  to  2,000  Ib.  After  filling,  the  cover  is  clamped  on,  first 
making  the  joint  with  some  flour-paste.  A  fire  of  wood  is  started 
under  the  retort,  low  at  first  and  increasing  to  a  cherry-red  heat 
at  the  end,  using  0.65  to  0.75  cord.  The  operation  lasts  10  to  14 
hours,  the  mercury  being  driven  off  down  to  I  to  1.5%.  Care 
must  be  taken  not  to  heat  the  retort  too  rapidly,  nor  to  raise 
the  heat  too  high  for  fear  of  blistering  it,  after  which  it  is 
allowed  to  cool,  and  the  silver  residue  removed.  The  melting 
down  of  these  residues  is  performed  in  plumbago  crucibles  with 
the  addition  of  a  little  soda  and  borax  as  fluxes.  When  melted, 
the  crucible  is  removed  from  the  furnace  by  means  of  basket 
tongs,  which  surround  and  grip  the  crucible  firmly  so  that  it  can 
be  lifted  out  by  the  chain  hoist  and  poured  into  an  ingot  mold  n 
in.  long  by  4.5  in.  wide  and  4.5  in.  deep,  and  capable  of  holding 
1 ,000  oz.  or  70  Ib.  of  silver. 

The  residue,  suspended  in  the  muddy  water  run  from  the 
settlers,  known  as  tailing,  contains  a  certain  portion  of  heavy 
unaltered  ore  which  can  be  generally  profitably  worked  either 
on  Frue  vanners  or  over  some  other  of  the  concentrating 
tables.  Otherwise  sluices  are  used,  two  or  three  together,  2  to  3 
in.  high,  20  in.  broad  and  as  much  as  1,800  ft.  long.  The  bottoms 
of  the  sluices  are  covered  with  coarse  blankets,  which  may  easily 
be  taken  up  and  washed  to  remove  the  concentrated  or  heavy 
portion  of  the  ore,  which  has  been  caught  in  the  meshes  of  the 
blankets. 

Cost  of  pan  amalgamation  (Washoe  process)  per  ton  of  ore 
treated : 

Power $0.087 

Labor    0-361 

Chemicals,  salt,  acid  and  bluestone 0.465 

Loss  of  mercury   0.750 

Wear  of  pans    0.200 

Wear  of  dies  and  shoes 0.400 

Oil,  interest  and  superintendence o.ioo 


Total  cost  per  ton $2.363 


OF    THE    COMMON    METALS.  185 

41.     THE  Boss  SYSTEM  OF  SILVER  MILLING. 

This  system  or  process  was  originated  by  M.  P.  Boss,  a  Cali- 
fornian  engineer,  and  differs  from  the  ordinary  settling-tank 
system  in  being  continuous,  saving  much  hard  labor.  It  may 
be  used  both  on  free-milling  ores  and  on  rebellious  ores  requiring 
a  preliminary  roast. 

The  ore-pulp  as  it  leaves  the  battery  passes  through  the  pans 
in  series  (see  Fig.  87  and  88),  being  finely  ground  in  the  first 
ones,  then  amalgamated  and  finally  settled,  the  pulp  being  ground, 


i      Cross  Section 

FIG.  87.      BOSS-PROCESS   SILVER   MILL    (ELEVATION). 

brought  into  intimate  contact  with  the  mercury  and  ultimately 
settled  in  series,  each  pan  doing  part  of  the  work.  The  first 
two  5-ft.  pans  of  the  series  do  the  fine  grinding  of  each  5 
stamps.  The  pulp,  which  is  a  thinner  mixture  of  ore  and  water 
than  in  the  tank  system,  flows  from  the  top  of  one  pan  by  a  short 
pipe  connection  to  the  next.  After  this  the  finely  ground  pulp 
passes  through  the  5-ft.  amalgamating  pans  and  then  through 
all  the  settlers  arranged  in  the  same  way  and  on  the  same  level. 
Each  pan  has  its  pipe  connection  to  the  next,  and  the  last  pan 


1 86  THE    METALLURGY 

in  the  same  way  to  the  settlers.  Any  pan  or  settler  can  be  by- 
passed or  cut  out,  the  pulp  flowing  around  to  the  following  ones. 
The  quicksilver  (as  well  as  the  salt  and  bluestone,  in  a  state 
of  solution)  is  fed  automatically  and  continuously  to  each  of 
the  amalgamating  pans,  the  portion  settling  to  the  bottom  col- 
lecting in  a  mercury-well.  The  settlers  also  have  mercury- 
wells,  so  that  the  collecting  of  the  mercury,  with  its  contained 
amalgam,  is  continuous.  As  it  accumulates  in  the  wells,  it  over- 
flows to  the  strainer,  where  the  excess  mercury  is  elevated  and 
returned  to  the  overhead  supply-tank.  The  continuous  discharge 
from  the  last  settler  goes  through  classifying-cones  to  concen- 
trating tables,  where  the  valuable  heavy  part  of  the  ore  is  re- 
moved as  concentrate. 

42.     THE  COMBINATION  PROCESS  OF  SILVER  MILLING. 

The  process  is  applicable  to  ores  carrying  silver  with  some 
gold,  and  containing  sulphides  of  base  metals,  such  as  pyrite, 
galena  and  zinc,  which  also  carry  gold  and  silver  values.  It  is 
necessary,  however,  that  the  silver  not  present  in  the  sulphides 
should  be  amalgamated. 

It  consists  in  wet  stamping  the  ore,  permitting  the  pulp  to 
run  over  amalgamated  plates  as  in  gold  milling,  concentrating 
out  the  heavy  gold  and  silver-bearing  sulphides,  pan  amalgamat- 
ing and  settling  the  tailing  from  the  concentrating  tables  as  in 
the  Washoe  process. 

When  compared  with  either  wet  or  roast  amalgamation  the 
process  has  much  to  recommend  it. 

By  wet  stamping  the  stamps  crush  one  and  one-half  to  two 
times  as  fast  as  dry.  It  is  true  that  the  roast  amalgamation  will 
extract  10%  more  than  can  be  done  by  raw  amalgamation,  but 
this  is  offset  by  the  cost  and  losses  of  roasting.  The  combination 
process  also  saves  the  lead  by  concentration,  and  removes  both 
the  base-metal  sulphides,  arsenic  and  lead,  which  tend  to  foul 
and  cause  loss  of  the  mercury  and  interfere  with  amalgamation ; 
and  there  results  a  cleaner  or  higher  grade  bullion.  Manganese 
minerals,  which  us-e  up  chemicals  in  the  pan,  are  also  removed 
by  concentration.  It  is  probable  that,  because  of  these  advan- 


FIG.  88.      BOSS-PROCESS  SILVER  MILL  (PLAN). 


1 88  THE    METALLURGY 

tages,  the  combination  process  will  extend,  certainly  at  the  ex- 
pense of  roast  amalgamation  which  is  fast  losing  ground,  and 
may  perhaps  supplant  lixiviation  either  by  hyposulphite  or  by 
cyanide. 

Fig.  89  is  a  perspective  view  of  a  lo-stamp  combination  mill. 
It  shows,  at  the  upper  stage,  a  car  about  to  be  dumped  over  a 
grizzly,  while  the  lumps  pass  on  to  the  crusher  where  they  are 
crushed  to  I  in.  diam.,  this  portion  then  joining  the  fine  in  the 
storage  bin  ready  for  the  stamps. 

Let  us  consider  an  oxidized  ore  containing  native  silver,  silver 
chloride,  a  little  silver  sulphide,  lead  and  copper,  and,  perhaps, 
mangenese  minerals  constituting  the  heavy  part,  and  a  light 
gangue  of  quartz,  clay  and  calcite.  The  problem  is  to  save  the 
gold  and  silver,  to  remove  the  lead,  copper,  arsenic  and  manga- 
nese. Arsenic  tends  to  sicken  the  mercury,  while  manganese 
uses  up  chemicals  in  the  pan  amalgamation. 

The  ore  is  fed  automatically  to  the  stamps,  each  of  which  may 
be  considered  as  crushing  5  tons  daily  to  pass  a  3O-mesh  screen. 
The  pulp,  issuing  from  the  mortar,  passes  over  an  apron-plate 
where  part  of  the  gold  and  the  silver  is  recovered.  Thence 
the  pulp  passes  to  concentrating  tables  where  the  heavy  part  is 
removed  and,  as  concentrate,  shipped  to  the  smelting  works.  The 
tailing  of  the  tables  flows  through  a  series  of  settling  boxes  as 
in  the  Washoe  process.  Indeed,  from  this  point  on,  operations 
are  conducted  as  in  that  process.  Bluestone  and  salt  are  used 
in  amalgamating  to  break  up  silver  sulphides.  It  may  be  noted 
that  the  careful  settling  of  the  tailing  ensures  the  eventual  de- 
livery to  the  pans  of  the  amalgam  and  mercury  from  the  apron- 
plates.  Five-foot  combination  amalgamating  pans  are  used 
which  treat  a  charge  of  1.25  tons,  the  time  of  treatment  varying 
from  4  to  8  hours,  according  to  the  nature  of  the  ore.  The  re- 
covery of  gold  and  silver  in  one  case  was  as  follows : 

Gold  %        Silver  % 

Recovered  on  the  apron-plates   22  3 

Recovered  on   concentrating  tables 28  32 

Recovered  in  amalgamating  pans 32  35 

Lost  in  tailing   18  30 

100  100 


OF    THE    COMMON    METALS. 


IQO  THE    METALLURGY 

Concentration  adds  but  little  to  the  costs  as  compared  with 
pan  amalgamation,  so  that  $3  per  ton  may  be  taken  as  a  fair 
figure. 

43.     CHLORIDIZIXG  ROASTING. 

This  is  performed  upon  rebellious  silver  ores  where  we  desire 
to  bring  the  contained  silver  into  the  form  of  a  chloride  so  that 
it  can  be  extracted  by  amalgamation  or  by  lixiviation.  Such 
ores  contain  the  silver  as  sulphide  or  as  arsenical  or  as  antimonial 
silver  sulphides  associated  with  the  heavy  sulphides  of  iron, 
zinc,  copper,  and  lead. 

As  a  preliminary  to  roasting  such  ores  are  dry  crushed,  either 
by  rolls,  or  by  stamps.  Galena-bearing  ores  are  preferably 
crushed  to  4O-mesh,  but  with  pyrite,  i6-mesh  is  sufficient. 

The  roasting  is  done  with  the  addition  of  salt,  and  in  presence 
of  3  to  8%  of  pyrite,  the  larger  quantity  being  needed  where 
there  is  much  limestone  in  the  ore.  If  the  ore  contain  more  than 
8%  pyrite  it  may  be  roasted  down  to  that  percentage  in  sulphur, 
after  which  the  salt  may  be  added.  The  amount  of  salt  varies 
from  3  to  18%  according  to  the  quantity  of  contained  pyrite, 
blende,  or  limestone,  all  of  which  take  up  chlorine. 

The  roasting  operation  is  at  first  an  oxidizing  one,  acting 
upon  the  heavy  metals,  and  converting  them  into  either  oxides 
or  sulphates.  It  may  be  divided  into  three  stages:  (i)  The 
kindling,  (2)  the  desulphurizing  or  oxidation,  and  (3)  the 
chlorination.  Referring  to  the  'Chemistry  of  Roasting,'  we  find 
an  account  of  what  occurs  in  the  kindling  stage.  We  find  that 
the  ferrous  sulphate,  which  has  been  formed  from  the  pyrite, 
begins  to  decompose  at  590°  C,  evolving  sulphuric  anhydride, 
and  this  reacts  upon  the  salt  thus  : 

2  Na  Cl+2  SO3=Na2  SO4-f  SCX+2  Cl 

2  X  97300  2  X  91800  328800  71000  =  +  2I600, 

an  exothermic  reaction.  The  chlorine,  thus  set  free,  begins  act- 
ing upon  the  silver  sulphides,  arsenides  and  antimonides,  con- 
verting the  silver  into  chloride. 

At  the  second  stage  of  the  roasting  •  above  mentioned,  es- 
pecially where  salt  has  not  been  at  first  added,  the  heat  may 


OF   THE    COMMON    METALS.  Ipl 

rise  to  such  a  degree  as  to  form  metallic  silver  according  to  the 

reaction : 

Ag2  S+02=2  Ag+S02 

3000  7 1 000=: + 68000, 

also  a  strong  exothermic  reaction.  In  the  subsequent  amalgama- 
tion, the  fact  that  the  silver  is  in  metallic  form  is  no  detriment,, 
but,  if  lixiviation  is  to  follow  roasting,  silver  is  but  little  acted 
upon  by  the  solution,  and  hence,  for  the  latter  process,  the  roast- 
ing must  be  more  cautiously  conducted. 

However,  metallic  silver  is  slowly  acted  upon  at  a  red  heat 
by  chlorine  gas  and  some  silver  chloride  is  formed. 

At  the  various  stages  of  roasting,  iron,  copper,  zinc  and  lead 
sulphates  are  formed,  which  are  acted  on  by  the  salt  as  illus- 
trated in  the  case  of  the  iron  sulphate, 

Fe  SO4+2  Na  Cl=Na2  SO4+Fe  C12,  so  that  soluble  chlorides 
of  iron,  copper  and  zinc  result,  besides  the  sparingly  soluble  Pb 
C12.  That  is  to  say,  the  salts  of  the  heavy  metals  are  brought 
in  soluble  form  as  the  result  of  a  chloridizing  roast.  These  de- 
compositions are  not,  however,  complete,  so  that  some  soluble 
sulphates  are  later  extracted  when  the  roasted  ore  is  treated 
with  water. 

For  moderate  quantities  of  ore  the  chloridizing  roast  is  carried 
on  in  reverberatory  roasters,  in  larger  quantities  in  one  or  an- 
other of  the  mechanical  roasters.  The  perfectly  dry  and  salted 
ore  is  charged  to  the  furnace  and  roasted  at  a  low  heat,  at  first 
to  kindle  the  ore  and  to  bring  about  decompositions.  The  heat 
is  raised  upon  it  toward  the  last  to  produce  the  chloridizing  re- 
actions. 

At  the  present  time  it  is  not  usual  to  continue  the  roasting  in 
the  furnace  with  a  view  of  converting  therein  all  the  silver  com- 
pounds into  chlorides,  but  the  charge  is  withdrawn  in  a  hot  con- 
dition before  this  stage  is  reached,  and,  during  the  gradual  cool- 
ing, lasting  12  to  30  hours,  a  further  chloridizing  takes  place,  due 
to  the  action  of  free  chlorine,  with  which  the  ore  is  saturated,  on 
yet  undecomposed  silver  sulphides.  This  may  increase  the  chlo- 
ridization  10  to  40%.  Upon  the  completion  of  this  operation  of 
heap  chlorination,  as  it  is  called,  and  upon  ores  containing  cop- 


192  THE    METALLURGY 

per  chloride,  a  wetting  down  or  sprinkling  will  result  in  an  addi- 
tional chlorination  of  3  to  6  %.  Thus,  at  the  Lexington  Mill, 
Butte,  Mont.,  the  ore,  just  after  roasting  in  a  Stetefeldt  furnace, 
was  chloridized  to  65%,  after  2  hours  to  75  to  80%,  and  at  the 
end  of  36  hours  to  92%  of  its  silver  contents. 

The  loss  of  silver  by  volatilization  when  the  ore  has  been  prop- 
erly roasted  should  not  exceed  8%,  except  in  presence  of  many 
volatile  elements  like  arsenic,  antimony,  selenium  or  tellurium. 
If,  however,  the  roasting  is  completed  at  too  high  a  temperature, 
the  volatilization  losses  may  run  up  to  18%. 

Taking  the  cost  of  a  chloridizing  roast  at  $2.40  per  ton,  it  will 
be  equal  to  the  value  of  4  oz.  at  6oc.  per  ounce.  Adding,  on  a 
25-oz.  ore,  8%  for  the  volatilization  loss,  or  2  oz.  more,  we  have 
a  total  of  6  oz.  or  24%,  so  that  60%  extraction  by  raw  amalgama- 
tion would  be  as  profitable  as  84%  on  roasted  ore,  to  say  nothing 
of  the  cost  of  installation  of  the  roasting  plant  and  the  extra  cost 
of  dry  crushing.  It  would  seeni,  therefore,  that  the  lower  grade 
silver  ores,  of  less  than  30  oz.,  could  be  more  profitably  treated 
raw. 

44.     DRY  SILVER  MILLING.     (REESE  RIVER  PROCESS.) 

This  process  is  applied  to  the  treatment  of  rebellious  silver  ores, 
in  which  the  silver  is  so  locked  up  that  it  is  required,  before  amal- 
gamation, to  roast  the  ore,  necessitating  dry  crushing.  While 
rebellious  ores  are  worked  by  this  method,  it  pays,  in  many  cases, 
where  smelting  works  are  available,  to  ship  the  ore  rather  than 
to  mill  it.  Or  again,  if  the  heavy  portion  of  the  ore  carries  the 
values,  it  may  be  profitable  to  undertake  concentration  only,  and 
to  submit  to  some  loss  in  the  tailing.  Such  problems  should  be 
worked  out  on  a  trial  scale  before  undertaking  the  construction 
of  a  plant. 

The  ores  mainly  worked  by  this  process  are  those  containing 
sulphides  of  silver,  particularly  antimonial  and  arsenical  sul- 
phides, together  with  base-metal  sulphides  containing  copper, 
iron,  zinc  and  lead.  The  latter,  however,  when  over  a  certain 
limit,  render  the  ore  unsuitable  for  this  process. 

The  ore  before  roasting  is  dry  crushed  either  by  rolls  or  by 
stamps.  The  former  method  is  fully  described  in  the  chapter  on 


194  THE    METALLURGY 

crushing;  the  ore  after  crushing  being  fed  continuously  to  the 
feed-hopper  of  the  roaster.  Otherwise  operations  are  carried  for- 
ward as  in  the  dry  crushing  silver  mill,  Fig.  90  and  91,  now  to 
be  described. 

The  ore,  as  in  wet  milling,  is  coarsely  crushed  during  the  10- 
hour  day  shift,  and  delivered  to  the  storage  bin,  whence  it  is 
continuously  fed  by  an  automatic  ore  feeder  during  the  24  hours 
to  a  cast-iron  revolving  dryer  18  ft.  long.  Here  the  ore 
is  dried  out;  and  by  inclined  launders  slides  to  the  automatic 
feeders  of  the  stamps,  entering  the  double  discharge  mortar 
where  it  is  finely  crushed.  As  fast  as  the  crushing  proceeds,  the 
ore  is  projected  by  its  impact  against  the  screen,  the  finer  particles 
passing  through,  while  the  coarser  material  drops  back  upon  the 
die  to  be  again  crushed.  Naturally  this  produces  dust,  and,  to 
prevent  much  of  it  in  the  mill,  the  mortar  is  housed.  There  is 
also  an  exhaust  .fan,  connected  to  the  housing,  by  which  the  dust 
is  carried  away  and  deposited  in  a  dust  chamber.  The  finely 
ground  ore,  passing  through  the  screens,  is  carried  by  screw- 
conveyors,  one  on  each  side  of  the  mortar,  to  an  elevator.  It  is 
customary  to  feed  to  the  last  five  stamps  the  rock  salt  necessary 
to  go  with  this  ore,  and,  which  being  finely  crushed,  can  be  in- 
timately mixed  with  it.  The  ore  and  salt  mixture  raised  by  the 
elevator  passes  by  screw-conveyors  to  the  feed-hopper  of  a 
\Yhite-Howell  roasting  furnace,  see  Fig.  40,  where  it  receives  a 
chloridizing  roast.  On  the  plan  view  (Fig.  91)  is  shown  the 
flue-chambers  which  collect  the  dust  and  the  stack  which  furnishes 
the  draft.  The  ore  from  the  roaster  falls  into  pits  or  receptacles 
where  it  undergoes  further  chlorination,  and  is  then  moistened  with 
a  little  water  to  prevent  dusting  and  placed  in  heaps  upon  the 
cooling  floor  to  be  treated  in  batches  or  charges  in  the  pans. 

The  pan  amalgamation  of  roasted  ores  is  performed  in  wooden- 
sided  pans  resembling  those  used  in  the  Washoe  process.  As 
compared  with  it,  the  yield  of  silver  is  greater,  being  as  high  as 
97%,  and  the  loss  of  quicksilver  less,  being  but  0.25  Ib.  per  ton  of 
ore.  Water  is  first  run  into  the  pan,  while  in  motion,  and  the  ore 
added,  in  1.5  ton  batches,  until  the  pulp  is  of  the  consistency  of 
honey.  If  the  ore  is  badly  roasted,  salt  and  bluestone  are  put  in 
to  decompose  silver  sulphide.  The  free  chlorine  in  the  ore  is 


OF   THE    COMMON    METALS. 


195 


196  THE    METALLURGY 

taken  up  by  the  iron  surfaces  of  the  pan,  and  from  any  iron  par- 
ticles, which  have  come  from  the  stamping,  and  is  thus  removed 
from  detrimental  action  on  the  mercury,  300  Ib.  of  which  is 
added  in  I  to  2  hours  after  starting.  The  muller  is  raised  at  the 
time  of  this  addition  so  that  no  further  grinding  takes  place,  the 
action  of  the  muller  being  to  mix  intimately  the  mercury  and  the 
pulp.  Amalgamation  proceeds  rapidly  at  first,  the  silver  chloride 
being  reduced  to  metal  as  in  the  Washoe  process.  The  iron  pres- 
ent reduces  the  higher  chlorides  of  copper  and  iron  and  converts 
any  mercurous  chloride  to  metal.  This  mixing  is  kept  up  for 
a  period  of  six  hours,  the  pulp  is  then  diluted  with  water,  and 
the  entire  charge  is  run  into  the  settler.  The  amalgam  is  col- 
lected and  treated  as  in  the  Washoe  process.  At  the  Lexington 
mill,  Butte,  Mont.,  an  ore  containing  28.5  oz.  Ag  and  0.58  oz. 
Au  per  ton  gave  a  yield  of  93.3%  of  the  silver  and  60%  of  the 
gold  after  roasting.  The  loss  of  silver  in  roasting  was,  however, 
7%,  while  the  loss  of  gold  was  20%. 

The  cost  of  dry  stamping,  followed  by  chloridizing  roasting, 
may  be  taken  at  $6.48,  being  the  average  of  three  mills  in  differ- 
ent parts  of  the  Rocky  Mountain  country,  which  vary  but  little 
from  one  another. 

45.     HYPOSULPHITE   LIXIVIATIOX    OF    SILVER    ORES.      (PATERA 

PROCESS.) 

Free-milling  silver  ores,  that  is,  oxidized  ores  containing  the 
silver  in  native  state  or  as  chloride,  do  not  need  to  be  treated  by 
this  method,  but  preferably  by  milling  and  amalgamation. 
Hyposulphite  lixiviation  is  used  upon  ores  containing  simple 
and  compound  sulphides  of  silver  which  have  undergone 
a  preliminary  chloridizing  roasting.  The  silver  is  then  dissolved 
by  sodium  hyposulphite,  precipitated  by  sodium  sulphide,  and  the 
precipitated  silver  sulphide  worked  up  for  the  resultant  silver. 
The  crystalized  sodium  hyposulphite  (Na2S2O3,  5  H,O),  contains 
5  equivalents  of  combined  water  and  100  parts  of  it  will  dissolve 
40  parts  of  silver  chloride.  Ores,  to  be  adapted  to  the  process, 
must  not  be  rich  in  lead,  copper  or  lime,  because  in  roasting,  their 
chlorides  tend  to  coat  over  the  undecomposed  particles  contain- 


OF    THE    COMMON    METALS.  197 

ing  silver.  They  also  go  into  solution  as  sulphates  and  chlorides. 
As  a  result,  we  find  the  lead  and  copper  precipitating  with  the 
silver,  thus  contaminating  the  silver  product,  while  the  quicklime 
produced  in  the  roasting  diminishes  the  solvent  power  of  the 
hyposulphite  and/with  the  lead  present,  precipitates  it  as  hydrate, 
coating  over  the  particles  of  silver  chloride  and  preventing  their 
lixiviation. 

Up  to  a  certain  point  this  may  be  prevented  by  a  preliminary 
extraction  of  these  soluble  base-metal  salts  by  hot  water.  Metal- 
lic silver,  if  finely  divided,  may  be  lixiviated,  as  also  the  arsenate 
and  antimonate  of  silver  produced  during  the  roasting.  Unde- 
composed  silver  sulphides  and  arsenic  or  antimony  sulphides  are, 
however,  not  attacked  by  the  hypo  solution. 

Roasting  the  ore.  The  ore  undergoes  a  chloridizing  roast,  as 
described  in  the  chapter  on  roasting,  it  being  much  the  same  as 
a  chloridizing  roast  for  amalgamation,  except  that  greater  care 
must  be  exercised,  the  coarser  unattacked  silver  being  only  slowly 
dissolved  by  hpyo  solution  while  it  is  readily  soluble  in  mercury. 

Leaching  the  ore.  To  give  an  idea  of  the  quantities  of  bases 
present  in  a  silver  ore  successfully  treated  by  lixiviation  we  will 
take  Ontario  ore,  Park  City,  Utah,  as  an  example.  The  silver 
sulphides  are  argentite  and  cerargyrite,  together  with  tetrahed- 
rite,  forming  'gray-silver  ore'  called  also  fahlerz.  Analysis 
gives  55.2%  SiO2,  13.1%  A12O3,  3.3%  Fe  and  Mn,  6%  Pb,  1.4% 
Cu,  9.6%  Zn,  1.4%  As  and  Sb,  7.7%  S.  This  ore  would,  there- 
fore give,  on  roasting,  the  soluble  sulphates  and  chlorides  of 
iron,  lead,  copper  and  zinc,  which  must  be  got  rid  of  before  treat- 
ing the  ore  with  hypo  solution  for  the  extraction  of  the  silver. 

Fig.  92  is  an  elevation  of  a  Hoffman  leaching  vat,  12  ft.  diam. 
and  4  ft.  deep,  with  a  central  discharge  opening  for  tailing.  The 
false  bottom,  as  shown,  slopes  toward  the  centre  and  is  a  wooden 
lattice  work  made  in  sections  and  covered  with  cocoa  matting. 
Upon  this  is  laid  a  canvas  filter-cloth,  6  in.  larger  in  diameter  than 
the  vat.  The  edges  of  the  canvas  are  crowded  down  by  means  of 
a  rope  driven  into  the  groove  p  at  the  inside  circumference  of  the 
vat,  and  in  the  same  way  at  pf  near  the  centre. 

The  ore  is  brought  from  the  cooling-floor  near  the  roasters  to 
the  tank,  and  tipped  in  evenly,  taking  care  that  it  is  not  tramped 


198  THE    METALLURGY 

down  or  compressed  in  any  way  to  interfere  with  its  loose,  porous 
condition.  It  fills  the  tank  to  a  depth  of  2.5  ft.,  this  thickness  be- 
ing preferred  because  the  water  used  in  the  first  wash  will  contain 
less  salt  (taken  from  the  ore)  to  leach  out  silver,  and  because  the 
contents  of  the  vat  are  more  easily  sluiced  out.  About  12-in.  space 
is  left  above  the  charge  to  the  top  of  the  tank. 

The  preliminary  leaching  with  water.  Cold  water  is  the  best 
for  leaching  where  there  are  but  small  quantities  of  the  base- 
metal  salts  to  remove ;  but  if  there  is  much  of  these,  then  hot  water 
must  be  used.  In  such  a  case  much  silver  chloride  is  dissolved, 
due  to  the  presence  of  salt  in  the  water  derived  from  the  ore.  The 
water  admitted  on  top  of  the  ore  should  leach  through  it  at  the 
rate  of  2  to  16  in.  per  hour  until  the  escaping  water  shows  no  pre- 
cipitate on  testing  with  sodium  sulphide.  The  quantity  of  wash- 
water  needed  per  ton  of  ore  is  I  to  2l/2  tons,  and  the  operation 
takes  from  2  to  24  hours.  If  the  wash- water  contains  but  little 
silver  chloride  it  is  run  to  waste.  If  much,  it  is  precipitated  in 
a  tank  to  which  it  has  been  run,  by  the  addition  of  sodium  sul- 
phide, which  will  precipitate  the  silver  as  sulphide,  together  with 
some  sulphides  of  base-metal  salts.  The  precipitate  from  this 
tank  is  later  treated  like  the  main  portion  of  silver  sulphide. 

Lixiviation  with  h\po  solution.  A  0.25  to  2.5%  solution  of  the 
thiosulphate  is  run  in  as  soon  as  the  wash-water  has  sunk  to  the 
level  of  the  ore,  forcing  the  latter  in  advance  out  of  the  vat,  while, 
as  soon  as  the  hypo  solution  follows,  it  is  diverted  to  the  precipi- 
tating tanks.  The  crystallized  hypo  salt  dissolves  silver  chloride 
according  to  the  following  formula : 

2  Xa2S203,  5  H20+2AgCl=AgS203,Na2S203+2NaCl+5H20; 
that  is,  the  crystallized  salt  dissolves  40%  of  its  weight  of  AgCl, 
or  30%  of  silver.  It  dissolves  not  only  AgCl,  but  also  some 
finely  divided  metallic  silver,  silver  oxide,  arsenate  and  antimonate 
and  gold.  Of  the  other  metals,  copper  dissolves  much  like  silver, 
while  lead  sulphate  and  calcium  sulphate  also  dissolve,  but  the 
solvent  properties  of  the  hypo  solution  are  diminished  in  presence 
of  lead  and  sodium  sulphates,  and  particularly  by  caustic  alkalies 
and  alkaline  earths,  such  as  quicklime,  the  latter  due  to  limestone 
which  may  have  existed  in  the  ore.  The  solvent  is  allowed  to 
leach  the  ore  until  the  silver  is  removed,  as  may  be  ascertained  by 


OF    THE    COMMON    METALS. 


199 


assaying  the  exhausted  ore  for  silver.  There  should  not  be  more 
than  4  oz.  per  ton  in  the  tailing,  the  extraction  amounting  to  70 
to  85%  of  the  silver.  The  operation  takes  from  6  to  53  hours. 
The  ore  is  then  treated  to  a  water-wash,  the  washings  also  going 
to  the  precipitation  tank,  10  ft.  diam.  by  Sy2  ft.  high. 

Precipitation  of  the  silver  out  of  the  hypo  solution.  This  is 
effected  by  the  addition  of  sodium  sulphide  to  the  solution  just 
in  sufficient  quantity  to  do  the  work,  the  sodium  hyposulphite  be- 


FIG.  92.     HOFFMAN  LEACHING  TANK. 

ing  regenerated  as  follows  :  Ag2S2O3,  Na2S2O3  +  Na2S  —  Ag2S 
+  2.  Na2S2O3.  Gold,  copper,  and  salts  of  the  base  metals  are  also 
thrown  down.  In  precipitating,  sodium  sulphide  is  added,  while 
the  solution  is  constantly  stirred,  until  a  sample  withdrawn  from 
the  vat  gives  but  a  very  slight  precipitate  with  sodium  sulphide. 
The  precipitate  is  allowed  to  settle  and  the  supernatant  clear 
solution  is  drawn  off,  the  precipitate  remaining  in  the  vat.  An- 
other charge  may  then  be  run  in,  and  treated  as  before.  The 
precipitate  thus  accumulates  in  the  bottom  of  the  tank.  At  the 
expiration  of  several  days  to  a  week  the  precipitate,  together  with 
the  solution  which  has  collected  at  the  bottom  of  the  tank,  is  run 


2OO  THE    METALLURGY 

to  a  filter-press,  where  the  solution  is  removed  and  sent  to  the 
stock-tank,  while  the  precipitate  is  taken  out  of  the  press  for 
further  treatment.  The  solution,  which  must  contain  no  excess  of 
sodium  sulphide,  is  then  returned  to  the  stock-tank  to  be  used 
again  on  other  charges  of  ore.  As  it  has  been  weakened  by  the 
addition  of  wash-water,  the  stock  solution  is  strengthened  by  the 
addition  of  fresh  hyposulphite. 

Sodium  sulphide  is  made  at  the  works  by  dissolving  caustic  soda 
in  its  own  weight  of  water  in  an  iron  kettle  at  80°  C,  and  adding 
gradually  to  it  powdered  sulphur.  The  addition  of  the  sulphur 
causes  the  liquid  to  swell  to  two  or  three  times  its  original  bulk, 
so  that,  to  begin  with,  the  pot'  should  be  no  more  than  one-quarter 
full.  The  sulphur,  equal  to  60%  of  the  caustic  soda,  dissolves  in 
a  few  minutes.  The  sodium  sulphide  thus  made  is  poured  into 
molds  where  it  solidifies.  For  use,  it  is  dissolved  in  water.  The 
amount  of  hypo  salt  consumed  is  2  to  4^2  lb.  per  ton  of  ore. 

Treatment  of  the  precipitated  sulphide.  After  filter-pressing, 
the  damp  precipitate  is  molded  into  cakes.  For  drying,  these 
are  placed  in  a  reverberatory  furnace,  16  by  6.5  ft.  hearth  dimen- 
sions, with  a  small  grate.  The  drying  must  be  done  at  a  mod- 
erate heat,  such  that  the  flame  shall  not  come  in  contact  with  or 
ignite  the  sulphides.  The  dried  precipitate  will  contain  18  to 
35%  of  silver.  Following  we  give  an  analysis  of  dried  sulphides 
from  the  Marsac  mill,  Park  City,  Utah : 

Ag          (11,360  oz.  per  ton) 34-7$% 

Au  ( 12.6  oz.  per  ton) 0.04 

Cu  21.60 

Pb  0.50 

Fe  0.75 

Sb  0.18 

ALO3    °-25 

Si62    0.25 

S 20.74 

This  dried  sulphide  is  generally  sold  to  smelting  works  where 
it  is  treated  as  follows,  to  obtain  the  silver: 

It  is  gradually  brought  upon  the  hearth  of  an  English  cupelling 
furnace  (see  Fig.  152),  where  it  is  roasted  and  the  silver  taken  up 


OF    THE    COMMON    METALS.  2OI 

by  the  molten  lead-bath.  The  other  metals  enter  the  litharge, 
which,  as  usual,  is  continuously  forming  and  flowing  from  such 
a  furnace.  The  litharge  is  sent  to  the  blast-furnace,  since  it  con- 
tains some  of  the  silver,  but  the  bulk  of  the  silver  goes  into  the 
lead-bath.  When  all  the  sulphide  has  been  treated,  the  lead  in 
the  test  or  hearth  is  cupelled  off,  leaving  the  molten  silver  ready 
for  molding  into  bars. 

The  extraction  or  yield  of  silver  by  the  Patera  process  is  70 
to  90%. 

At  Sombrerete  and  at  Cusihuiriachic,  Mex.,  the  Russell  process 
has  been  given  up  for  the  Patera  process.  At  the  former  place 
the  ores  contain  galena  (9  to  10%  Pb)  blende,  copper  and  iron 
pyrite,  and  a  silicious  gangue  together  with  silver  sulphides,  the 
ore  having  41.9  oz.  Ag  per  ton.  The  ore  loses  in  roasting  4.8% 
and  the  extraction,  figured  on  the  raw  ore,  is  82.5%. 

The  cost  of  treatment  per  ton  is : 

Crushing   $i  .36 

Roasting  and  salt  (6% ) 2.68 

Labor  in  leaching 0.27 

Chemicals    0.30 

Superintendence    i  .02 

Heating,  lighting,  pumping  and  repairs 0.08 


Cost  per  ton $5-7I 


46.     THE  RUSSELL  PROCESS. 

The  Russell  process  is  a  modification  of  the  Patera  process. 
This  modification  consists  principally  in  the  use  of  another  solu- 
tion in  addition  to  the  hypo  solution  used  in  the  latter  for  the 
extraction  of  the  silver.  By  mixing  in  solution  two  parts  of  the 
hypo  salt  with  one  part  of  copper  sulphate,  we  obtain  a  double 
salt,  Na2S2O3,  Cu2S2O3,  called  the  extra  solution,  which  has 
nine  times  the  solvent  power  for  native  silver,  for  silver  sul- 
phides, and  for  silver  arsenides  and  antimonides  that  the  or- 
dinary hypo  solution  has.  In  the  case  of  an  imperfectly  roasted 
ore,  the  use  of  the  extra  solution  ensures  the  extraction  of  an 


2O2  THE    METALLURGY 

additional  amount  of  silver  from  the  just-named  compounds, 
which  could  not  have  been  gotten  out  by  the  ordinary  solution. 

Referring  to  the  diagram,  Fig.  93,  and  to  the  practice  at  the 
Marsac  mill,  Park  City,  Utah,  the  process  is  conducted  as  fol- 
lows:  The  ore  contains  76.6%  SiO2,  1.65%  Fe,  1.32%  CaO, 
0.23%  MgO,  3.5%  Pb,  0.39%  Cu,  5.3%  Zn,  0.7%  S  with  37.3  oz. 
Ag  and  0.049  oz-  ^u  Per  ton.  The  ore,  mixed  with  8.9%  salt, 
is  stamped  dry  through  a  3O-mesh  screen,  each  stamp  having 
a  capacity  of  2.33  tons  daily.  The  ore  is  submitted  to  a  chlor- 
idizing  roast  in  a  Stetefeldt  furnace,  having  a  capacity  of  70 
tons  daily,  and  able  to  convert  92.4%  of  the  silver  into  chloride. 

After  roasting,  the  ore  is  charged  into  a  leaching  tank  such  as 
already  described  in  the  Patera  process,  17  ft.  diam.  by  9  ft. 
deep  and  capable  of  holding  72  tons.  In  this  tank  it  is  treated 
with  a  wash  of  water,  using  18  cu.  ft.  per  ton  of  ore,  the  water 
percolating  through  it  at  the  rate  of  nearly  4  in.  per  hour.  This 
solution  passes  then  to  the  base-metal  precipitating  tank,  where 
the  base  metals  were  thrown  down  as  sulphides  by  the  addition 
of  sodium  sulphide.  The  precipitate  contained  on  an  average 
4.2%  Pb,  3.9%  Cu,  4,800  oz.  Ag,  and  2.7  oz.  Au  per  ton,  and 
was  sold  to  the  smelter. 

The  ore  is  next  treated  with  a  1.5%  hypo  solution,  by  which  a 
greater  part  of  the  silver  is  extracted,  the  solution  going  to  the 
lead-precipitating  tank.  This  work  takes  87  hours.  Now  comes 
the  'extra'  solution  consisting  of  0.75%  copper  sulphate,  CuSO4, 
5  H2O  and  2.25%  of  the  crystallized  hyposulphite  Na2S2O3, 
5  H2O,  by  which  a  further  amount  of  silver  is  taken  out,  being 
those  undecomposed  sulphides  unattacked  by  the  ordinary  solu- 
tion. This  operation  takes  27  hours,  the  filtrate  being  also  run 
to  the  lead-precipitating  tank.  In  this  tank  lead  carbonate,  insol- 
uble in  hypo  solution,  is  precipitated,  using  5  Ib.  of  sodium  car- 
bonate per  ton  of  ore  treated,  according  to  the  reaction: 
Na,  CO3+Pb  C12=2  Na  Cl+Pb  CO3.  This  precipitate  contains, 
on  an  average,  32%  Pb,  1,000  oz.  Ag  and  1.2  oz.  Au  per  ton. 

After  settling,  the  supernatant  liquor  is  transferred  to  the 
silver  precipitating  tank,  and  the  residue  in  the  tank  is  sent  to 
the  filter-press.  The  precipitated  lead  carbonate  is  sold  to  the 
smelting  works,  while  the  filtrate  is  stored,  to  be  sent  later  to 


OF   THE    COMMON    METALS. 


203 


the  silver  precipitating  tank.  The  solution  in  the  silver  pre- 
cipitating tank  is  now  treated  with  just  enough  sodium  sulphide 
to  bring  down  the  silver  as  in  the  Patera  process.  The 
precipitated  silver  sulphide  contains  on  an  average  25%  Cu, 


~anoex  or  cre 

Isr.  Preparation  of  tfie  Mater/mis 
2nd.  L/xiyiatton  or  extraction  ofttuMeteta 
3rd.  PmcipSf&fion  of  thefletats 
4th.  Treatment  af  Product 


FIG.  93.      FLOW-SHEET  FOR  RUSSELL  PROCESS. 

40%  or  11,700  oz.  Ag,  and  n.6  oz.  Au  per  ton.  It  is  transferred 
to  the  filter-press  where  the  solution  is  removed,  and,  together 
with  that  from  the  tank,  is  returned  to  the  stock-solution  tank 
to  be  used  over  again.  Thus  the  filter-presses  and  dryer  take 
care  of  the  base-metal  precipitates,  the  silver  precipitate  and  the 
precipitated  lead  carbonate.  The  precipitate  is  dried  and  sacked, 
being  then  sold  to  the  smelter,  and  worked  up  by  them  in  the 


2O4  THE    METALLURGY 

English  cupelling  furnace  as  already  described  (see  Fig.  155). 
The  costs  by  the  Russell  process,  based  upon  an  output  of  100 
tons  ore  daily,  may  be  given  as  follows : 

Crushing  and  roasting $4.62 

Labor  in  leaching 0.83 

Tools,  light,  pumping  and  heating 0.12 

Chemicals    0.92 

Repairs  and  superintendence 0.18 

Assaying   0.08 

Treatment  of  out-products o.io 

$6^5 
The  extraction  is  85^,  the  proportion  of  silver  obtained  being, 

In  base-metal    sulphides 6.2% 

In  lead   carbonates 2.6 

In  silver    precipitates 75.7 

In  mill   cleanings 0.5 

85.0 

The  Russell  process  has  proved  successful  in  exceptional  cases 
only.  We  have  already  alluded  to  its  having  been  given  up  in 
two  places  in  Mexico,  while  elsewhere  failures  have  occurred. 
Compared  with  the  Patera  process  the  cost  of  the  chemicals  is 
greater  (92c.  against  42c.),  the  plant  is  more  complicated,  and 
greater  skill  is  needed  to  work  it.  It  can,  of  course,  be  applied 
to  raw  oxidized  ores  or  to  those  which  have  an  ordinary  dead 
roast,  and  although  the  yield  of  silver  is  greater  when  the  extra 
solution  is  used,  yet  this  is  offset  by  the  consumption  of  copper 
sulphate.  In  presence  of  much  galena  and  zinc  blende,  the  extra 
solution  seems  to  extract  the  silver  no  better  than  the  ordinary 
solution.  Nor  does  the  process  work  well  in  presence  of  much 
lime,  being  too  slow  and  consuming  much  copper  sulphate. 

47.     THE  LEACH IXG  OF  ARGENTIFEROUS  MATTES. 

In  order  to  extract  the  silver  from  such  mattes,  which  also 
carry  silver,  two  different  processes  have  been  used,  the  Augustin 
and  the  Ziervogel.  In  both  methods  the  principle  consists  in 


OF    THE    COMMON    METALS. 


205 


bringing  the  silver  in  soluble  form,  either  as  the  chloride  or  the 
sulphate,  and  then  leaching  it  out  with  a  suitable  solvent,  the 
silver  being  precipitated  in  presence  of  metallic  copper. 

The  Augustin  process. — The  matte  is  crushed,  roasted  with 
salt  to  form  silver  chloride,  leached  with  brine  to  extract  the 
silver,  the  silver  in  the  brine  solution  precipitated  on  plates  of 


PRELIMINARY 
CALCINATION/          A 
&  SMELTING1        ** 


.    5ASTINS 
8      .'WITH  SALT. 


FIG.  94.    FLOW-SHEET  FOR  AUGUSTIN  PROCESS. 

copper,  and  the  copper,  in  the  clear  decanted  solution,  precipi- 
tated by  scrap  iron. 

The  Ziervogel  process. — At  the  Boston  and  Colorado  smelting 
works  at  Argo,  Colo.,  the  process  is  used  upon  copper  mattes  com- 
paratively free  from  arsenic,  antimony  or  bismuth,  all  of  which 
tend  to  form  insoluble  compounds  with  silver.  It  is  performed 
as  follows : 

The  matte  is  made  in  reverberatory  furnaces  from  ores  con- 
taining gold,  silver  and  copper  and  is  cast  in  sandbeds.  Its  com- 
position is  47.3%  Cu,  8.1%  Pb,  2.7%  Zn,  17.7%  Fe,  21.6%  S 
with  400  oz.  of  Ag  and  10  oz.  Au  per  ton.  It  is  crushed  and  rolled 
to  6-mesh,  and  then  sent  to  a  Pearce  turret  roasting  furnace  where 
it  receives  a  preliminary  roasting  at  a  low  heat,  removing  the 
sulphur  down  to  6.3%,  and  converting  iron  and  copper  to  oxides 


2O6  THE    METALLURGY 

and  sulphates  (see  Chemistry  of  Roasting).  The  partly  roasted 
matte  then  goes  to  Chilean  mills  where  it  is  finely  ground  to 
6o-mesh. 

It  is  now  treated  in  charges  of  1,600  lb.,  to  a  sulphatizing  roast 
in  small  single-hearth  reverberatory  furnaces.  In  the  process, 
the  remaining  sulphides  are  converted  to  sulphates,  and,  at  a 
little  higher  heat,  the  copper  sulphate  begins  to  decompose  so 
that  no  more  than  i%  is  finally  left.  In  so  doing  the  sulphuric 
anhydride  evolved  acting  powerfully  on  metallic  silver,  transforms 
it  to  the  state  of  sulphate.  It  has  been  found  that  the  addition 
of  2%  sodium  sulphate  (salt  cake)  greatly  accelerates  this 
change.  The  roasting  occurs  in  four  stages : 

At  the  first  stage,  with  the  draft  checked  and  the  working 
doors  open,  the  charge  is  kept  at  a  low  temperature  for  1.5  hours. 
It  becomes  heated  evenly  throughout  and  glows  from  the 
oxidation  of  Cu3S  to  Cu2O.  At  the  second  stage,  during  1.5 
hours,  the  heat  is  slightly  increased  while  the  charge  is  con- 
stantly rabbled.  Sulphate  of  iron  is  decomposed  and  CuSO4 
is  formed.  The  charge  swells,  and  becomes  porous  and  spongy 
from  the  formation  of  this  salt.  In  the  third  stage,  for  an  hour 
the  temperature  is  again  increased,  until  tests  show  that  the 
silver  is  'out,'  that  is,  in  the  condition  of  sulphate  soluble  in 
water.  During  this  stage  the  copper  sulphate,  decomposing, 
transmits  its  SO3  to  the  silver  compounds,  forming  sulphate.  At 
the  fourth  stage,  with  the  temperature  constant,  the  charge  is 
gathered  together  and  bruised  down  with  a  heavy  paddle  to 
break  up  the  lumps,  and  it  is  then  vigorously  stirred  to  oxidize 
any  remaining  Cu2O  to  CuO,  and  to  decompose  nearly  all  the 
copper  sulphate.  This  temperature  is  not  further  increased,  since 
it  would  have  the  effect  of  decomposing  the  silver  sulphate  to 
Ag2O,  thus  again  making  it  insoluble. 

The  progress  of  the  roasting  is  tested  by  putting  small  samples 
from  time  to  time  in  hot  water  and  bringing  the  soluble  sulphates 
in  solution.  Early  in  the  third  stage  it  becomes  deep  blue;  and 
later,  as  the  silver  sulphate  begins  to  form,  it  is  immediately  re- 
duced to  silver  spangles  by  the  cuprous  oxide  present.  As  the 
stage  proceeds,  and  the  copper  sulphate  is  decomposed,  the 
solution  becomes  less  blue  and  the  silver  spangles  increase  and 


OF   THE    COMMON    METALS. 


207 


then  diminish.  During  the  fourth  stage  the  Cu2O  is  changed  to 
CuO,  and  the  spangles  disappear,  while  a  light  blue  color  still 
remains,  due  to  a  little  copper  sulphate  still  remaining,  which 
indicates  that  the  silver  sulphate  it  not  itself  becoming  decom- 
posed. 

(A  sample  thus  roasted,  but  at  another  works,  and  containing 
246  oz.  silver,  showed  by  analysis  2.5%  Fe  SO4  +  Zn  SO4;  0.6% 


[FINAL  LIQUOR, 


•ARKET  COPPER. 

FIG.  95.     FLOW-SHEET  FOR  ZIERVOGEL  PROCESS. 

Cu  SO4,  and  i.o  AgSO4,  or  of  silver  213.9  oz-  Per  ton-  There  was 
left,  therefore,  in  the  ore  32.1  oz.  or  13%  of  the  silver  in  insoluble 
form.) 

Leaching. — This  is  performed  in  tubs  or  tanks,  each  holding 
1,000  Ib.  of  matte,  and  furnished  with  filter  bottoms.  The  cop- 
per-bearing solution  goes  through  a  series  of  vats  containing 
copper  plates  upon  which  the  silver  precipitates.  The  silver-free 
solution  containing  copper  then  goes  to  tanks  containing  scrap 
iron  where  it  is  precipitated. 

The  cement  silver  from  the  precipitating  vat  is  next  transferred 
in  charges  of  3,000  oz.  to  a  tub  where  dilute  sulphuric  acid  ( i  to 
100)  is  added,  and  it  is  boiled  by  forcing  in  a  mixture  of  air  and 
steam  from  an  injector.  The  air  oxidizes  the  copper,  still  re- 
tained in  the  precipitate,  to  CuO  which  is  dissolved  by  the  acid, 


2O8  THE    METALLURGY 

while  the  steam  keeps  the  solution  boiling.  In  two  or  three  hours 
the  whole  of  the  copper  has  been  dissolved.  The  copper  sulphate 
solution  is  drawn  off,  and  the  precipitate  repeatedly  washed  with 
hot  water  until  quite  free  from  copper.  It  is  then  transferred  to, 
and  dried,  in  a  long  pan  set  over  a  coal  fire,  and,  after  drying, 
melted  down  in  a  wind  furnace  into  ingots  of  999  to  999.5  fine. 

Residues  from  the  leaching  tanks  or  tubs. — These,  still  con- 
taining one-tenth  of  their  original  silver,  say  40  oz.  per  ton, 
all  the  gold  (10  oz.  per  ton)  and  the  copper  as  CuO,  are  sent 
to  a  reverberatory  furnace  where,  together  with  rich  ore  both 
silicious  and  sulphide,  they  are  melted  together  to  a  matte.  The 
slag  from  the  treatment  goes  back  to  the  ore-smelting  furnace, 
while  the  matte,  which  has  been  tapped  into  sand  molds,  is  sent 
to  another  reverberatory  furnace  to  be  treated  according  to  the 
English  method  of  making  'Best  selected  copper.'  Here  the  pigs 
of  matte  are  piled  up  near  the  bridge  of  the  furnace  and  are 
slowly  roasted  and  melted  down  under  the  oxidizing  action  of 
the  flame  and  air.  The  copper  oxide  formed  in  the  'roasting/  re- 
acting upon  the  matte,  forms  copper  according  to  the  reaction. 

2  Cu2O+Cu2S=6  Cu+SO2. 

It  is  aimed  to  carry  on  the  oxidation  of  the  copper,  so  that  the 
amount  formed  will  give  one  part  of  copper  to  fifteen  of 
matte.  The  matte  is  now  tapped  from  the  furnace  into  sand- 
molds,  and,  in  the  first  molds,  will  be  found  the  plates  of  copper 
overlaid  with  the  matte.  These  bottoms  have  carried  down  with 
them  the  impurities,  as  arsenic,  antimony,  lead  and  bismuth, 
together  with  practically  all  the  gold  (100  to  200  oz.),  and  some 
of  the  silver.  On  the  other  hand,  the  matte  has  risen  to  the  stage 
of  white  metal  of  75%  copper,  still  holding  90  to  100  oz.  Ag  per 
ton,  but  not  more  than  0.2  oz.  of  gold,  and  is  sent  to  be  again 
treated  by  the  Ziervogel  method,  its  roasting,  however,  being  done 
in  separate  furnaces  from  that  of  the  first-matte.  The  residues, 
after  this  second  treatment,  are  principally  copper  oxide  contain- 
ing but  10  oz.  Ag  per  ton,  and  are  sold  to  the  oil  refiners.  At 
the  Argo  works,  the  bottoms,  heretofore  treated  for  the  extrac- 
tion of  the  precious  metals  by  a  secret  process,  could  be  more 
profitably  treated  by  electrolytic  refining. 


OF  THE  COMMON  METALS.  2CK) 

48.     CYANIDING  SILVER  ORES. 

Silver  ores  carrying  some  gold  and  in  which  the  silver  occurs 
not  only  as  native  silver,  but  as  chloride,  argentite  and  stephanite, 
have  been  more  or  less  successfully  cyanided.  Thus  at  Chloride 
Point,  Utah,  where  the  silver  occurs  as  chloride,  the  extraction 
averages  71%;  at  the  Palmarejo  mine,  Chihuahua,  Mex.,  the 
extraction  is  54%  of  the  silver  and  96%  of  the  gold;  and,  at 
El  Salvador,  Mex.,  85  to  90%  of  the  silver  and  90  to  92%  of 
the  gold  is  recovered. 

The  main  points  in  cyaniding  silver  ores  are  these : 

i. — That  a  comparatively  long  time  is  needed,  usually  10  to  25 
days,  since  silver  compounds  are  more  difficultly  soluble  than  gold. 

2. — Thorough  oxygenation  is  necessary,  due,  not  only  to  the 
larger  quantity  of  silver  to  be  dissolved,  but  also  to  the  fact  that 
the  silver  compounds  need  to  have,  at  least,  initial  oxidation  to 
make  them  more  soluble  in  cyanide  solution.  To  insure  this, 
double  treatment  is  necessary,  that  is,  the  ore,  after  having  been 
leached  in  one  vat,  is  shoveled  out  into  another  where  the  treat- 
ment is  repeated.  In  so  doing,  it  becomes  thoroughly  exposed 
to  the  air.  Also,  if  the  solution  is  allowed  to  sink  several  inches 
below  the  top  of  the  charge  before  fresh  solution  is  run  on,  the 
air  is  drawn  down  and  permeates  the  ore.  In  the  slime  treat- 
ment, of  course,  the  pulp  may  receive  thorough  aeration  by  agita- 
tion with  air. 

3. — Stronger  solutions  must  be  used  than  in  the  treatment  of 
gold  ores.  The  first,  or  strong  solution,  may  be  0.7%,  and  the 
weak  solution  0.25%,  while,  for  gold,  a  0.25%  solution  would  be 
the  strong,  and  0.05%  a  suitable  weak  solution. 

4. — The  consumption  of  potassium  cyanide  is  higher  than  in 
the  case  of  gold  ores.  It  varies  from  2.5  to  4  Ib.  per  ton  com- 
pared with  0.4  to  0.8  Ib.,  consumed  for  gold  ores. 

5. — The  precipitation  of  silver  by  zinc  shavings  presents  no 
difficulty  and  is  practically  complete;  and  despite  the  fact  that 
relatively  so  much  silver  is  precipitated  as  compared  with  gold,  no 
more  zinc  is  used  up. 


PART  V.    IRON 


PART  V.     IRON. 

49.     PIG  IRON — ITS  MANUFACTURE  IN  THE  BLAST-FURNACE. 

The  three  kinds  of  ores  used  in  making  pig  iron  are  carbonates, 
magnetites  and  hematites. 

(i)  Carbonate  ore  (FeCO3),  spathic  iron,  black-band,  clay- 
band  or  clay  iron-stone  contains  48.3%  Fe  when  pure.  It  is  often 
roasted  to  drive  off  the  CO2  before  going  to  the  blast-furnace.  In 
England  it  is  the  principal  ore  used,  but  in  the  United  States, 
though  widely  distributed,  it  is  too  low  grade  to  be  used  in  com- 
petition with  richer  ores.  (2)  Magnetite  (Fe3O4)  contains 
72.4%  Fe  when  pure.  It  is  a  valuable  source  of  iron,  there  being 
deposits  of  it  in  New  York,  Pennsylvania,  New  Jersey  and 
Michigan.  The  New  York  beds  occur  in  the  Lake  Champlain 
district,  and  on  account  of  the  presence  of  titanium  have  been 
considered  valueless,  it  having  been  asserted  that  the  titanium 
produces  an  infusible  slag.  This,  however,  has  been  proved  an 
error,  and  in  future  the  metallurgist  need  not  be  deterred  by 
such  a  fear  from  using  titaniferous  ores.  In  Pennsylvania  the 
Cornwall  deposits  are  the  most  important,  and  can  give  a  pig 
iron  of  not  more  than  0.04%  P,  but  the  ore  (containing  2.5%  S) 
should  first  be  roasted.  The  copper  in  the  ore  is  the  most  objec- 
tionable element  and  will  be  found  in  the  pig  to  the  amount  of 
0.5  to  0.75%.  This  is  unobjectionable  in  the  finished  material, 
but  it  produces  hot-shortness  and  many  imperfections  in  the 
rolling.  In  New  Jersey  occur  extensive  beds  of  low-grade  sili- 
cious  hematite  which  have  been  utilized  by  Edison  in  the  pro- 
duction of  iron  ore  on  a  commercial  scale.  He  has  mined, 
crushed  and  concentrated  the  ore  and  made  it  into  briquettes 
which  contain  as  little  as  3.3%  SiO2  0.04%  P,  and  as  high  as  67% 
Fe.  It  is  doubtful,  however,  whether  the  enterprise  can  be  car- 
ried on  at  a  profit  in  competition  with  foreign  ores.  (3)  Hema- 
tite (Fe2O3)  contains,  when  pure,  70%  Fe.  It  occurs  as  specular 


214  THE    METALLURGY 

hematite,  having  a  black,  hard,  shining  look,  or  combined  with 
more  or  less  water,  having  a  reddish-brown  or  yellowish-brown 
color.  It  is  customary  to  speak  of  the  higher  grade  ores  as  red 
or  brown  hematites,  and  of  the  more  hydrous  ones,  as  soft  hema- 
tites or  limonites.  Strictly  speaking  a  limonite  is  a  bog-iron  ore 
containing  over  20%  H2O.  However,  the  other  nomenclature 
has  the  sanction  of  custom. 

50.     IRON  BLAST-FURNACE. 

Fig.  96  is  a  sectional  elevation  of  a  blast-furnace  in  which 
coke  is  used  as  a  fuel.  Beginning  at  the  bottom  we  have  a  heavy 
foundation  (i)  upon  which  is  carried  the  hearth  (15)  and 
the  columns  (4)  which  support  the  upper  brick  work  con- 
stituting the  shaft  of  the  furnace.  The  hearth  (n  ft.  diam. 
by  9.5  ft.  high)  to  contain  the  molten  iron  and  slag,  ex- 
tends from  the  foundation  to  just  above  the  tuyeres  (22).  It 
has  a  bottom  and  sides  of  fire-brick,  and,  at  its  lowest  interior 
point,  an  iron  tap  (27)  whence  the  molten  pig  iron  is  withdrawn 
when  a  quantity  has  accumulated.  At  (24)  is  shown  the  cinder 
notch  by  which  the  slag  or  cinder  is  tapped  off.  The  crucible  is 
surrounded  by  a  hearth- jacket  (26)  of  steel,  cooled  on  the  out- 
side by  sprays  of  water  playing  against  it,  and  thus  withstanding 
the  corrosive  action  of  the  molten  slag  upon  the  brickwork.  Air, 
under  a  pressure  of  5  to  14  Ib.  to  the  square  inch,  enters  at  the 
tuyeres  (21),  and  care  must  be  taken  to  withdraw  the  slag  before 
it  can  rise  to  this  level.  This  air  is  supplied  by  the  tuyere-stock 
(33)  from  the  bustle-pipe  (13),  which  surrounds  the  furnace  and 
connects  with  the  main  supply  pipe,  which  brings  the  air  to  the 
furnace.  Proceeding  upward  from  the  hearth,  the  furnace  widens 
out  for  22  ft.  at  the  level  of  the  mantel  (5),  the  top  of  the  sup- 
porting columns,  where  it  is  23  ft.  diam.,  this  expanding  portion 
being  called  the  bosh.  In  this  region  or  zone  occurs  the  forma- 
tion of  slag,  so  that  the  brickwork  of  the  bosh  is  subject  to  a 
slagging  or  scouring  action.  To  prevent  this,  water-cooled  bosh- 
plates  (14)  are  laid  into  the  brickwork  all  around  the  furnace 
at  every  18  in.  to  2  ft.  apart  vertically.  Beginning  from  the  top 
of  the  bosh  and  extending  to  the  throat  of  the  furnace  (9)  we 


OF    THE    COMMON    METALS. 


215 


FIG.  96.     IRON    BLAST-FURNACE. 


2l6  THE    METALLURGY 

have  the  shaft.  This  is  supported  by  the  mantel  beams  which  rest 
on  the  columns  (4). 

The  upper  part  of  the  furnace  is  closed  by  a  bell  (47),  and  has 
at  the  side  an  outlet  for  the  escape  of  the  smoke.  The  inwall 
(69)  of  the  stack  is  of  fire-brick,  while  the  main  portion  of  it  is 
built  of  common  brick.  It  is  sheathed  with  a  steel  shell  (46). 
When  in  operation,  the  furnace  is  filled  to  just  below  the  outlet 
part  of  this  level,  it  being  known  as  the  stock-line,  and  here  the 
furnace  is  15  ft.  diam.  As  the  stock  sinks  charges  are  put  in,  to 
keep  the  stock-line  at  this  level.  The  total  height  of  the  furnace 
is  80  feet. 

Ordinarily  the  top  of  the  furnace  is  kept  closed  by  the  conical 
bell  (47),  which  is  suspended  from  the  ends  of  the  counter- 
weighted  beams  (55).  .The  bell  closes  the  bottom  of  a  circular 
hopper  (48)  into  which  the  charge  is  dumped.  At  the  moment 
of  charging,  the  outer  end  of  the  lever  is  raised  by  a  piston 
working  in  an  air  cylinder  (60),  the  bell  is  lowered,  and  the 
charge  slides  into  the  furnace,  and  the  bell  is  again  closed.  In 
this  interval  gas  and  flame  escape  from  the  furnace,  but,  in 
general,  the  gas  passes  off  by  the  down-comer  (39)  to  the  dust- 
catcher  (40)  and  thence  by  the  goose  neck  (41)  to  the  under- 
ground flue  which  leads  to  the  stoves  and  boilers,  where  the  gas 
is  burned.  On  top  of  the  down-comer  stands  the  bleeder  (37), 
by  which  the  gas  may  be  let  off  when  it  is  desired  to  relieve  the 
top  pressure  due  to  the  escaping  gases.  It  is  occasionally  used. 

At  many  furnaces,  the  stock  or  materials  of  the  charge  is 
raised  in  hand-barrows  or  buggies  to  the  furnace  top  or  tunnel- 
head  (51)  by  means  of  vertical  platform  hoists.  A  more  recent 
arrangement  consists  in  the  use  of  a  mechanical  charging  ap- 
paratus shown  in  Fig.  97.  A  double  bell  is  here  used,  by  which 
the  escape  of  the  gases  is  avoided.  The  charge  is  dropped  from 
the  skip  into  the  upper  oval-shaped  hopper  where  it  is  retained 
until  the  lower  hopper  is  empty.  The  smaller  upper  bell  is  now 
emptied  into  the  lower  hopper,  and  is  closed  to  receive  another 
charge.  The  charge  in  the  lower  hopper  is  dropped  when  needed 
by  lowering  the  large  lower  bell.  It  slides  largely  to  the  walls,  the 
stock  at  the  middle  being  a  little  lower  and  coarser  than  at  the 
sides.  The  charging  of  the  skip  (of  which  there  are  two,  run  in 


OF   THE    COMMON    METALS. 


FIG.  97.     IRON   BLAST-FURNACE  WITH   AUTOMATIC   CHARGING. 


218 


THE    METALLURGY 


balance)  is  thus  performed.  Iron  ore,  limestone  and  coke  come 
in  by  an  overhead  track  and  are  delivered  to  sloping-bottom 
storage  bins.  A  charge-car,  electrically  driven  and  with  a  weigh- 
ing attachment,  can  be  brought  to  any  bin  and  there  receive  a 
weighed  amount  of  ore  or  coke.  The  loaded  car  is  then  trans- 


FIG.  98.     COWPER  HOT-BLAST  STOVE  (ELEVATION). 

ferred  to,  and  discharged  into,  the  skip.  In  case  of  accident  to 
the  charge-car,  or  any  other  trouble  at  these  bins,  the  furnace  can 
still  be  kept  going  by  the  use  of  hand-barrows  or  buggies,  taking 
the  stock  from  piles  which  have  been  made  beneath  the  track 
at  the  right.  Hoisting  is  effected  by  a  hoisting  engine  set  well 
out  of  the  way  at  the  top  of  the  stock-house. 


OF    THE    COMMON    METALS. 


51.     THE  STOVES. 


219 


For  the  most  efficient  operation  of  an  iron  blast-furnace  it  is 
necessary  that  the  air  blown  into  it  should  be  heated,  generally 
from  500  to  750°  C.  To  accomplish  this,  a  furnace  is  equipped 
with  three  or  four  regenerative  fire-brick  stoves,  say  80  ft.  high 
and  14  ft.  diam.,  of  which  Fig.  98  and  99,  called  the  Cowper 
stove,  is  an  example.  It  consists  of  a  tight  steel  shell,  lined  with 
fire-brick  and  containing  a  checkerwork  of  specially  shaped  brick, 
having  frequent  openings  or  passages  from  top  to  bottom  of  the 
stove.  The  gases  from  the  blast-furnace,  containing,  say  24% 


FIG.  99.    COWPER  HOT-BLAST  STOVE  (PLAN). 

CO,  flowing  along  an  underground  flue  from  the  goose-neck  of 
the  dust-catcher  enter  at  g,  while  air  is  admitted  at  a,  the  two 
burning  in  the  flue  /,  and  heating  it  and  the  checkerwork  in  their 
passage  downward  to  the  valve  s,  and  thence,  by  an  underground- 
flue  to  a  tall  stack  200  ft.  high.  The  stove,  having  been  heated, 
the  valves  at  gf  a  and  s  are  closed  and  the  cold-air  valve  C  ad- 
mits air  under  pressure  from  the  blowing  engine,  the  hot  air 
valve  d,  at  the  bottom  of  the  combustion  chamber,  being  at  the 
same  time  opened.  The  air,  flowing  through  the  checkerwork, 
becomes  highly  heated.  Meanwhile  the  gas  has  been  turned  into 
the  other  stoves,  and  is  heating  them  up  in  the  same  way.  When 
the  first  stove  has  received  the  blast  for  half  an  hour,  it  is  turned 
into  the  next  just-heated  stove  and  so  on.  The  extended  surface 
of  the  checkerwork  serves  to  absorb  a  large  amount  of  heat  from 


22O  THE    METALLURGY 

the  burning  gases,  and  to  give  it  up  subsequently  to  the  blast. 
It  will  be  noticed  that  the  air  enters  at  the  coolest  and  leaves  at 
the  hottest  part  of  the  stove,  thus  insuring  the  maximum  rise  in 
temperature.  In  the  course  of  the  half  hour  the  hot  air  has 
dropped  100°  C  in  temperature,  at  least. 

52.     BLAST-FURXACE  PLANT. 

Fig.  100  represents  a  modern  blast-furnace  in  process  of  con- 
struction, showing  the  forked  down-comer  with  its  bleeder,  and 
the  upper  portion  of  the  dust-catcher ;  also  the  inclined  hoist. 
There  are  four  stoves,  each  as  high  as  the  furnace  itself.  Imme- 
diately back  of  the  stoves  is  to  be  seen  their  stack.  At  the  right 
of  the  picture  may  be  seen  the  steel  frame  of  the  double  vertical 
hoist  or  elevator  of  an  older  and  lower  furnace,  together  with  two 
of  the  stoves  belonging  to  it. 

53.     OPERATION  OF  THE  FURNACE. 

The  ore  is  dumped  into  the  furnace  with  50  to  65%  of  its 
weight  of  coke,  and  the  limestone  needed  to  form  a  predetermined 
slag.  The  furnace  should  be  at"  least  65  ft.  high,  and  is  now 
commonly  built  80  to  100  ft.  high.  It  is  kept  full  of  stock,  and 
the  combustion  of  the  coke  is  kept  up  by  the  introduction  of  the 
blast  at  a  high  pressure  at  the  tuyeres.  The  constant  tapping  off 
of  the  liquid  products  (slag  and  iron),  and  the  constant  addition 
of  fresh  stock,  prevent  the  heat  from  spreading  upward  as  it 
might  otherwise  do. 

The  actual  melting  zone  extends  several  feet  above  the  tuyeres. 
At  about  4  ft.  from  the  tuyere  opening,  where  it  enters  the  fur- 
nace, the  most  intense  combustion  occurs,  and  here  with  abundant 
access  of  air,  the  coke  gets  first  burned  to  CO2  while  in  the 
immediately  adjacent  zone  the  CO2  may  be  regarded  as  being 
changed  to  CO.  This  latter  operation  dissolves  carbon  out  of  the 
coke,  thus  lessening  its  quantity ;  and  the  reaction,  being  endother- 
mic,  lessens  the  intensity  of  the  heat.  The  upward-passing  gases 
are  CO  and  X.  Now,  we  must  remember  that  the  FeO,  reduced 
above  from  Fe2O3  ,  and  even  unreduced  ore  where  it  has  run  the 


OF    THE    COMMON    METALS. 


221 


FIG.  100.      IRON   BLAST-FURNACE. 


222  THE    METALLURGY 

gauntlet  of  the  CO,  are  brought  directly  in  contact  with  the 
glowing  coke,  and  reduction  is  completed. 

The  CO,  rising  through  the  furnace,  reduces  FeO  to  Fe  or 
spongy  iron.  On  reaching  the  lower  part  of  the  furnace,  the 
spongy  iron  with  carbon  becomes  pig  iron,  the  car- 
bon forming  a  carbide  of  iron  containing  3  to  4%  C.  The 
iron  also  absorbs  silica,  phosphorus  and  sulphur  from  the  earthy 
ingredients  of  the  charge. 

From  the  point  of  view  of  the  ideal  metallurgical  process  we 
thus  have,  in  the  blast-furnace : 

Zone  of  Preparation.     CO.,  and  moisture  driven  off. 

Zone  of  Reduction.     The  iron  reduced  to  FeO  and  then  to  Fe. 

Zone  of  Fusion.  Slag  formed;  the  iron,  as  a  carrier  or  col- 
lector, taking  up  C ;  Si ;  P ;  and  S. 

In  addition  to  the  coke  and  the  ore  there  is  limestone.  In  the 
upper  third  of  the  furnace  the  CO2  of  the  limestone  is  expelled, 
leaving  CaO.  The  action  is  endothermic  and  has  its  advantage 
in  lowering  the  temperature  of  the  upper  part  of  the  furnace. 
The  quantity  of  limestone  to  make  a  proper  slag,  and  lime  enough 
to  remove  S,  is  calculated  in  advance. 

The  gases  from  the  furnace  are  about  as  follows : 

f       CO  =24% 

Temperature    260°    C .         CO2=i6% 

(       N     =60% 

and  are  used  for  burning  in  stoves  and  under  the  boilers.  For- 
merly the  air  was  heated  in  iron  pipe  stoves,  but  today  the  regen- 
erative stoves  of  brick  are  used. 

In  the  foregoing  descriptions  it  has  been  assumed  that  coke 
is  the  fuel  employed,  and  this  is  true  in  the  majority  of  cases. 
Anthracite  coal  has  been  used  for  economic  reasons,  but  even 
then  more  satisfactory  practice  is  attained  when  coke  forms  part 
of  the  charge.  Such  furnaces  are  called  anthracite  furnaces,  but 
such  a  name  is  somewhat  misleading.  Some  furnaces  use  char- 
coal exclusively,  since  this  fuel  is  supposed  to  give  an  iron  of 
great  toughness,  particularly  valuable  in  making  cast-iron  car 
wheels,  and  other  articles  where  toughness  is  required.  Its 
superiority  over  other  irons  of  similar  chemical  constitution  has 


OF    THE    COMMON    METALS.  223 

been  widely  disputed,  but  there  is  much  testimony  in  favor  of 
charcoal-iron. 

Twenty-five  years  ago  the  practice  of  blast-furnaces  was 
regulated  by  rule-of-thumb  methods,  "  born  of  a  bigoted 
belief  on  the  part  of  ignorant  furnace-men  that  particular  ores 
and  fuel  could  be  worked  in  a  furnace  only  on  certain  special 
lines,  and  that  it  was  impious  to  drive  a  furnace  faster  than 
a  certain  rate  established  by  time-worn  tradition."  In  1879  cer- 
tain experimental  work,  instituted  at  the  Edgar  Thompson  steel 
works,  Pittsburg,  showed  that  it  was  possible  to  increase  the 
output  of  a  furnace  enormously  by  increasing  the  air  supply.  It 
was  also  found  that  the  amount  of  air,  not  the  pressure,  deter- 
mined the  speed. 

Under  this  new  system  it  was  thought  necessary  to  make  a 
steep  angle  bosh  of  80°,  but,  with  the  more  rapid  driving,  reduc- 
tion fell  off;  or  rather,  to  get  reduction,  the  fuel  had  to  be  kept 
high  (one  ton  of  coke  to  one  ton  of  pig  iron)  and,  where  coke 
was  expensive,  this  was  a  serious  matter.  Mr.  E.  C.  Potter,  of 
Chicago,  however,  showed  that  by  reducing  the  bosh  angle  to 
75°  and  by  using  somewhat  less  blast,  it  was  possible  to  cut  the 
coke  consumption  considerably  (from  2,240  down  to  1,800  or  to 
1,750  lb.).  Larger  furnaces  were  then  built  with  a  larger  hearth, 
which  also  increased  capacity. 

54.     MANAGEMENT  OF  THE  IRON  BLAST-FURNACE. 

Blowing  in. — The  furnace  is  first  dried  out  by  a  wood  fire  in 
the  crucible  for  several  days.  The  lower  part,  half  way  up  the 
bosh,  is  filled  with  cord-wood,  then  follows  a  heavy  bed  of  coke 
with  some  limestone  to  flux  the  coke  ash.  This  is  followed  by 
successive  layers  of  the  normal  charge  of  coke  with  charges  of 
gradually  increasing  amounts  of  ore  and  limestone  and  de- 
creasing amounts  of  slag,  until,  at  last,  the  normal  charge 
of  ore  and  flux  has  been  reached.  The  furnace  is  then  lit  at  the 
tuyeres,  and  a  weak  blast  put  on,  the  pressure  being  gradually 
increased  for  24  hours  until  the  regular  pressure  is  reached. 

Regular  running. — Ore  flux  and  fuel  (called  collectively 
'stock')  is  brought  in  buggies  from  the  stock-house  to  the  scales, 


224  THE    METALLURGY 

weighed,  hoisted  to  the  top  of  the  furnace  and  dumped  automat- 
ically into  the  hopper,  or  wheeled  by  'top-fillers'  to  the  bell  and 
dumped.  Fuel  is  dropped  by  itself,  the  ore  and  stone  (flux)  to- 
gether. The  level  of  the  charge  is  kept  just  below  the  outlet  for 
the  down-corner.  The  temperature  of  the  blast  (550  to  750°  C) 
is  kept  constant,  as  well  as  the  volume  of  the  blast,  by  regulating 
the  revolutions  of  the  engine.  The  slag  is  tapped,  say  every  two 
hours,  the  iron  every  four  to  six  hours.  If  the  metal  flows  from 
the  runner  into  the  molds  or  depressions  of  the  sand-bed,  it  is 
cooled  down  by  water,  and  while  hot  broken  by  sledges,  trans- 
ferred to  cars  or  trucks  and  stacked  in  the  pig-metal  yard  until 
cold.  Here  it  is  broken  and  assorted  according  to  fracture  or  by 
analysis.  If  the  metal  is  collected  in  ladles,  a  switching  engine 
takes  it  to  the  casting  machine. 

Blowing  down. — When  a  furnace  is  to  be  put  out  of  blast, 
charges  are  stopped  and  a  layer  of  coke  put  in.  The  materials 
gradually  descend,  and  this  is  kept  up  as  long  as  the  iron  and 
cinder  will  flow  out,  the  blast  being  gradually  diminished,  until 
finally  it  is  stopped  altogether  and  the  remaining  contents  of  the 
furnace  withdrawn  by  a  hole  broken  through  the  brick  work  at 
the  bottom. 

55.     IRON  SMELTING. 

The  air  entering  the  tuyeres  need  be  sufficient  to  burn  the  fuel 
only  to  CO,  according  to  the  reaction : 

C  +  O  =  CO, 

giving  29,000  Ib.  calories  or  2,415  per  Ib.  of  carbon  and  using  up 

O  X  12.38 

—    =72  cu.  ft.  air,  or  per  pound  of  coke  of  85%  carbon, 

0-^3  X  ^ 

61  cu.  ft.  of  air.  The  usual  way  of  measuring  the  amount  of  air 
that  enters  the  furnace  is  to  calculate  the  cubic  displacement  of  the 
pistons  of  the  blowing  engines,  but  this  is  not  accurate,  since  it 
makes  no  allowance  for  leakage  and  imperfect  action  of  the  valves, 
or  for  clearance.  With  this  allowance  we  find  65  to  70  cu.  ft. 
of  free  air  needed  per  pound  of  fuel.  On  Lake  ores,  where  1,600 
to  2,000  Ib.  of  coke  are  needed  per  long  ton  of  pig  iron,  this  will 
be  112,000  to  140,000  cu.  ft.  of  air,  or,  for  a  furnace  of  400  tons 


OF   THE    COMMON    METALS.  225 

capacity,  31,000  to  39,000  cu.  ft.  per  minute.  In  rising  through 
the  furnace,  part  of  the  CO  is  oxidized  to  CO2  by  reaction  on  the 
ore  ;  and  the  limestone  evolves  more,  so  that  the  tunnel-head  gases, 
with  a  consumption  of  1,900  Ib.  per  ton  of  pig  may  have  a  com- 
position 14.5%  CO2,  27.0%  CO  and  58.5%  N,  or  a  ratio  of 


C02 

Of  this  gas,  one-third  is  used  to  heat  the  stoves,  the  remainder 
being  mostly  utilized  for  producing  steam  at  the  boiler  for  the 
blowing  engines.  Since  the  efficiency  of  the  engine  is  no  more 
than  10%  of  the  total  heat  delivered  to  it,  the  gas-engine,  using 
the  furnace  gas,  and  having  an  efficiency  of  30%,  has  been  pro- 
posed in  place  of  the  steam-engine.  This  would  leave  a  surplus 
of  gas  available  for  giving  power  for  other  purposes  about  the 
works,  but  the  initial  cost  of  installing  such  a  power  plant,  and 
mechanical  difficulties  in  its  operation,  have  heretofore  stood  in 
the  way  of  introducing  it  to  any  great  extent. 

56.     THE  CHEMICAL  REACTIONS  OF  THE  IRON  BLAST-FURNACE. 

A  blast-furnace  may  be  regarded  as  an  immense  gas-producer, 
in  which  there  is  a  column  of  coke  (mixed  with  the  iron  ore  and 
limestone)  70  ft.  high,  varying  in  temperature  from  a  white-heat 
at  the  tuyeres  to  a  black  heat  at  the  tunnel-head.  As  soon  as  the 
hot  air  of  the  blast,  entering  at  the  tuyeres,  strikes  the  white-hot 
coke  there  is  an  immediate  formation  of  CO2,  followed  by  its  in- 
stantaneous reduction  to  CO.  Fig.  101  shows  graphically  the 
amount  and  progress  of  the  reactions,  while  the  temperatures 
and  heights,  at  which  the  reactions  take  place,  are  shown  at  the 
section  of  the  furnace  at  the  extreme  left  of  the  diagram.  To  pro- 
duce a  ton  of  pig  iron  (2,240  Ib.)  there  is  to  be  used  3,520  Ib.  of 
60%  iron-ore  containing  3,020  Ib.  Fe2O3,  1,388  Ib.  of  coke  and 
1,010  Ib.  limestone.  At  the  tunnel-head,  the  ore  (Fe2O3)  is 
plunged  into  an  atmosphere  containing  24%  CO,  16%  CO2  and 
60%  N  at  a  temperature  of  260°  C  ;  and  there  is  immediately  a 
reduction  of  part  of  the  ore  Fe,O3  according  to  the  slightly 
exothermic  reaction. 

3  Fe203  +  CO       =       2  Fe304  +  CO2 


226  THE    METALLURGY 

3  X  199.400  29,000  2  X  265,800  97,000=  -f  11,400  Cal.  ; 
this  reaction  increasing,  and  being  completed  by  the  time  the  ore 
has  reached  a  depth  of  10  ft.  and  at  a  temperature  of  450°  C  (see 
section  2).  During  this  period  begins  that  peculiar  reaction  of 
carbon-deposition,  by  which  the  gases  react  on  the  ore  and  de- 
posit carbon  throughout  its  pores.  This  mixture  descends  through 
the  furnace  and  appears  in  the  pig  iron  to  the  extent  of  say  4% 
of  its  weight. 

Continuing  its  descent,  by  the  time  it  has  reached  a  depth  of 
19  ft.  and  a  temperature  of  580°  C,  the  just  formed  Fe3O4  be- 
comes changed  to  FeO  (  see  column  3  )  as  follows  : 

Fe3O4  +  CO  =  3  FeO  +  CO2 

265,800  29,000  3X66,400  97,000=  -{-i,  400  Cal. 

The  FeO,  thus  formed,  impregnated  with  carbon,  descends  with 
little  change  until,  at  a  temperature  of  700°  C,  the  CO  gas  reacts 
upon  it  and  spongy  iron  begins  to  form,  the  reaction  being  com- 
plete when  the  temperature  has  reached  800°  C  and  at  a  depth 
of  32  ft. 

The  reaction  is  thus  expressed  : 

FeO  +  CO  =  Fe  -f  CO2 

66,400  29,000         97,000  =+  i,  600  cal. 

At  this  same  temperature,  the  limestone,  accompanying  the  iron 
ore,  by  the  expulsion  of  its  combined  CO,,  has  been  converted 
into  quicklime,  and  this  latter,  at  the  zone  of  fusion,  will  unite 
with  and  flux  the  silicions  portion  of  the  charge.  From  this 
depth  of  29  ft.  just  mentioned  and  to  the  depth  of  19  ft.  or  550° 
C,  the  CO.,  evolved  from  the  limestone,  reacting  on  the  coke,  dis- 
solves some  of  it  according  to  the  reaction 


CO,  +  C  = 
97,000         2  X  29,000  =  —  39,000  cal. 

absorbing  heat  from  the  charge  and  consuming  coke.  However, 
the  coke  remains  but  little  changed  until  it  reaches  the  region 
of  the  tuyeres,  as  is  shown  in  column  (6)  of  the  diagram  Fig.  101. 
From  the  point  of  32  ft.  and  of  800°  C  reactions  have  practi- 
cally ceased,  the  chief  one  now  being  a  reduction  of  a  small 


OF    THE    COMMON    METALS. 


227 


amount  of  FeO,  left  decomposed  by  the  CO.     This  is  gradually 
reduced  as  follows : 

FeO  -f  C  =  Fe  +  CO 
66,400  29,000  =  —  37400 

The  temperature  now  rises  gradually  and  uniformly  until  the 
intense  combustion  at  the  tuyeres  brings  it  to  a  maximum  of  at 
least  1,500°  C.  Of  the  two  gases  which  enter  the  furnace,  nitro- 
gen comprising  77%  of  the  whole  is  by  far  the  largest  constitu- 

(1)         (2)         (3)        (4)      (5)        (6)     (7)    (8)    (9)      (10)        (11)        (12)  (13) 


m 


13' 


32L 


36 


m 


(tee 


56' 


£00 


I 


£:£  «      3  • 

FIG.  101.      CHEMICAL   REACTIONS   OF   BLAST-FURNACE. 


ent,  or  indeed  the  largest  single  ingredient  column  as  shown  by 
12  of  the  diagram. 

Studying  closely  columns  10  and  n,  we  may  note  that  the  CO0, 
formed  so  freely  at  the  tuyeres,  is  at  once  changed  to  CO,  the 
glowing  coke  being  dissolved  by  the  CO2.  The  carbon  monoxide 
rises  unchanged,  until  it  begins  to  act  on  the  iron  oxides  at  the 
32-ft.  level  with  the  formation  of  CO2.  This,  united  to  that 
evolved  from  the  limestone,  forms  the  total  of  escaping  CO0  gas. 

It  must  be  observed  that  the  temperatures  and  heights  are  here 
given  as  a  definite  problem,  but  the  quantities  vary  with  every 
furnace,  and  even  in  the  same  furnace. 


228  THE    METALLURGY 

^ 

57.     CALCULATION   OF   IRON-BLAST   FURNACE  CHARGE. 

The  accompanying  charge-sheet  illustrates  the  method  of  esti- 
mating a  charge  for  an  iron  blast-furnace,  producing  a  pig  iron 
and  a  slag  of  a  certain  assumed  composition.  It  is  desired  to 
smelt  lumpy  Maxwell  ore  with  the  fine  Mesaba,  which,  alone, 
would  give  too  compact  a  charge  for  smelting.  The  pure  Mesaba 
would  make  too  little  slag  to  absorb  the  sulphur;  hence  it  is 
better  to  mix  with  the  Mesaba  a  certain  amount  of  the  silicious 

CHARGE    SHEET 


Name  of  Ore 

Weight 

SiO2 

Fe 

CaO 
&MgO 

s. 

A1203 

H20 

Wet 

1060 
1030 

Dry 

1000 
1000 

% 

Wt. 

% 

4.0 
18.0 

Wt. 

% 

so.o 

50.0 

Wt. 

<Tc 
1.0 

Wt. 

10 

or 
% 

;, 

Wt. 
8 

Mesaba  No.  1552 
Maxwell  No.  27 

6.0 
3.0 

40 
180 

600 
500 

1.0     10 
2.8     28 

Limestone 

" 

600 

600 

3.0 

18 

53.0 

318 

0.8      5 

Coke 

" 

1000 

4.4 

44 

1.0 

1.0 

10     1.2     12 

3.5     35 

282 

1100 

338 

20 

78 

For  pig   = 

65 

2SL 

slag  = 

217 

u 

PIG 

Fe  =  92.3% 
C  =    3.4 

S=    2.5  or  SiO2  =  5.4 
P=    1.0 
S=    0.2 


SLAG 

SiOo  =34.0% 
AloO3  =  14.0 
Fe'O  =  0.3 
CaO  =45.0 
MgO  =  4.0 
S  =  1.4 

Maxwell  ore.  The  high  content  of  lime  here  prescribed  also  en- 
sures elimination  of  the  sulphur.  We  therefore  specify  equal 
weights  of  each  ore,  and  note  down  their  percentage  values  in  the 
appropriate  columns.  Assuming  that  we  are  to  count  on  2,000  Ib. 
coke  per  ton  of  pig,  we  will  need  for  the  1,100  Ib.  of  iron  ore 
shown  in  the  column  (Fe)  about  1,000  pounds  of  coke,  which  we 
accordingly  specify.  This  coke  carries  1 1  %  of  ash  analyzing  40% 
SiO2,  9%  CaO  and  32%  A12O3. 

The  corresponding  SiCX  in  the  coke  will  therefore  be  11%  of 
40% ,  or  4.4% ,  and  the  percentages  for  Fe  and  CaO,  found  in  the 
same  way,  are  written  in.  The  sulphur  is  reported  as  a  percent- 
age of  the  coke  itself.  The  limestone  or  dolomite  (magnesia  be- 


OF    THE    COMMON    METALS.  229 

ing  calculated  as  below  specified)  is  the  variable  which  must  be 
made  to  conform  to  the  rest  of  the  charge,  and  its  amount  is  found 
by  a  cut-and-try  method  as  follows :  We  put  down  what  we 
judge  to  be  the  proper  amount,  say  600  Ib.  Write  down  now  all 
percentage  amounts  and  carry  out  the  weights  of  each  constitu- 
ent in  pounds,  summing  them  up  as  shown.  Referring  to  the 
analysis  of  the  pig  iron  we  will  have  1,100  Ib.  of  Fe,  or  nearly 
1,200  Ib.  of  pig  containing  5.4%  SiO2,  or  (neglecting  fractions) 
65  Ib.,  and  this,  deducted  from  the  total,  leaves  217  Ib.  to  enter 
the  slag.  Magnesia  is  a  stronger  base  than  lime  in  the  ratio  of 
their  atomic  weights,  or  as  40  to  56,  so  that  the  percentage  of 
MgO  multiplied  by  1.4  will  give  us  the  equivalent  lime,  in  this 
case  5.6%.  This,  together  with  the  45%,  also  given,  makes  a 
total  of  50.6%,  or,  in  ratio  to  the  silica  of  the  slag,  as  34  to  50.6 
or  as  i  to  1.49.  For  217  Ib.  of  SiO2  on  the  charge  sheet  we  there- 
fore require  327  Ib.  of  CaO  (too  much  by  n  Ib.),  or,  since  lime- 
stone is  half  CaO,  we  will  reduce  the  quantity  by  22  Ib.,  or,  in 
round  numbers,  20  Ib.  Erase  the  600  and  the  corresponding 
figures  in  the  other  columns,  writing  in  the  amended  amounts, 
when  it  will  be  found  that  the  calculation  will  be  correct  to  within 
the  limits  of  error  in  weighing.  Indeed,  variations  of  20  Ib.  in  the 
constituents  may  easily  occur,  due  to  this  cause.  Summing  up 
we  have  for  the  slag: 

SiCX  =  217  Ib.  =  34.0% 
A1,O3  =    78  "  =14.0     (12.4%  actual) 
CaO  +  MgO  =327_."  =50.6 

625  98.6 

Referring  to  the  slag  desired,  three  constituents,  namely,  SiO2, 
A12O3  and  CaO  (MgO)  will  amount  to  98.6%.  The  actual 
amount  of  A12O3  is,  however,  78  X  98.6 

-=  12.4% 
625 

Since  alumina  is  indifferent,  being  in  solution  in  the  slag  rather 
than  an  active  base,  we  will  find  that  the  slag  will  readjust  itself, 
so  that  the  percentages  of  SiO2  and  CaO  will  increase  enough  to 
make  up  for  the  deficiency,  i.e.  from  97.0  to  98.6%,  or  by  1.7% 
gives 


230 


THE    METALLURGY 


SiO,  =  34.0     34-6 

A1,O3  =  12.4     12.6 

CaO  =  50.6     51.4 


97.0    98.6 

In  deciding  on  what  slag  to  use,  we  may  choose  from  the  fol- 
lowing table  one  which  will  give  most  nearly  the  proportion  suited 
to  the  composition  of  the  iron  ores  used. 


SiQ2    I   FeO   |    CaO     |  MgO  |  A12Q3  |CaS 


Middlesborough,  England 
Saulines     France 

30.0 
31.6 

0.7 
0.6 

32.8 
47.2 

5.3 
1.4 

28.0 
17.0 

1.9 

Cuban  ore 

332 

40.7 

11.1 

13.7 

Lake  ore 

355 

405 

8.9 

12.0 

347 

1  3 

401 

109 

11.3 

Cuban  ore  

34.5 

46.5 

7.9 

10.5 

From  this  table  we  may  select  a  slag  which  shall  approximate 
the  most  nearly  in  A12O3  to  that  produced  by  the  ores  of  the 
charge. 

58.     CASTING  THE  PIG  IROX. 

Every  four  to  six  hours  the  iron  is  tapped  from  the  furnace, 
being  say  50  tons  at  a  time.  The  floor  of  the  cast-house,  40  ft. 
wide  and  80  to  150  ft.  long,  is  a  sand-bed  into  which  depressions 
are  molded,  connected  by  a  main  channel  or  'runner,'  which  re- 
ceives the  stream  of  molten  iron  flowing  from  the  furnace.  These 
molds  are  successively  filled,  forming  the  pigs  of  iron  weighing 
loo  Ib.  each. 

To  reduce  the  cost  of  handling  the  pig  metal,  casting  machines 
have  been  introduced  at  modern  plants.  Fig.  102  is  a  Uehling 
casting  machine  consisting  of  two  parallel  endless-chain  convey- 
ors, composed  of  molds,  each  capable  of  holding  120  Ib.  of  iron. 
The  iron  is  tapped  from  the  blast-furnace  into  a  large  ladle, 
shown  at  the  right  of  the  picture  and  thence  into 
the  near  end  of  the  conveyors,  and  as  it  passes  along, 
is  partly  cooled  by  water-sprays  so  that,  by  the  time  it  has 
reached  the  end  of  the  conveyor,  it  is  solid  and  can  be  dumped 
on  another  conveyor  having  a  flat  top.  Here  it  is  further  cooled 
by  the  conveyor  dipping  into  a  water  trough,  so  that  the  pigs  fall 
into  cars  quite  cool. 


OF    THE    COMMON    METALS.  23! 

59.     DISPOSAL  OF  CINDER. 

When  operating  a  charcoal  furnace,  having  an  output  of  100 
tons  of  pig  per  day,  so  little  cinder  is  made  that  it  is  often  cus- 
tomary to  permit  it  to  flow  out  after  the  iron  as  the  latter  is 
tapped,  setting  a  plate  across  the  stream  to  skim  off  the  slag,  and 
throwing  it  into  sand  beds,  from  which,  when  cooled,  it  is  carted 
away. 

A  large  coke  blast-furnace,  yielding  500  tons  of  pig  daily,  is 
quite  different.  It  produces  in  that  time  some  300  tons  of  cinder, 


FIG.  102.     UEHLING  CASTING  MACHINE. 

which  is  taken  off  at  the  cinder-tap  and  carried  by  launder  to  an 
adjoining  sunken  track,  where  stands  a  1 4-ton  cinder  car  to 
receive  it.  This  car  is  hauled  out  by  a  locomotive  to  the  dump, 
sometimes  a  mile  away,  where  the  contents  of  the  ladle  are  poured 
out  at  the  side  of  the  track.  The  track  is  gradually  raised  as  the 
dump  grows. 

60.     PIG  IRON. 

The  iron  produced  in  the  blast-furnace  is  not  pure,  but  con- 
tains 3.5  to  4.0%  of  carbon.  Some  of  this  carbon  is  chemically 
combined,  and  some  held  in  suspension  as  graphite.  If  a  large 
proportion  is  combined,  the  fracture  of  the  iron  looks  white  and 
the  metal  is  hard  and  brittle.  If  a  large  proportion  is  in  the  free 
state,  the  fracture  will  be  gray  or  black,  with  scales  of  graphite, 


232 


THE    METALLURGY 


and  the  iron  is  soft  and  tough.  Iron  also  contains  silica,  taken  up 
in  the  furnace  from  the  ash  of  the  coke.  It  varies  in  quantity 
from  0.5%  to  sometimes  as  much  as  3%,  but  is  usually  of  i  to 
2%.  A  small  proportion  of  sulphur  will  also  be  present,  seldom 
less  than  0.02%,  and  often  0.25%,  or  even  more.  With  over 
0.10%,  the  iron  is  apt  to  be  hard  and  brittle.  The  percentage  of 
silicon  and  sulphur  depends,  in  large  measure,  upon  furnace  con- 
ditions, and  hence  can  be  controlled,  but  there  is  one  element 
which  cannot  be  controlled  when  present,  and  which  goes  onto 
the  pig,  namely  phosphorus.  In  irons,  used  in  making  steel,  ac- 
cording to  the  usual  or  acid  bessemer  process,  the  phosphorus  is 
not  allowed  to  exceed  0.10%,  or,  in  the  ore,  no  more  than  0.5%. 
This  ensures  toughness  in  castings  made  from  it.  On  the  other 
hand,  phosphorus  confers  great  fluidity  on  cast  iron,  and  such 
irons,  containing  as  high  as  3.0  %  P,  are  in  demand  where  intri- 
cate castings  are  to  be  made,  or  where  brittleness  is  of  less  im- 
portance. 

Cast  iron  has  the  following  physical  characteristics  :  ( I )  It  is 
more  fusible  than  wrought  iron.  (2)  It  is  brittle.  (3)  It  cannot 
be  forged  either  hot  or  cold. 

Pig  iron  is  graded  according  to  its  appearance  into  several 
classes  or  grades,  ranging  from  the  soft  gray  or  No.  I  iron  to 
the  hard  white  or  No.  6  grade  as  follows : 

Alabama  Pig:  Iron. 


Grade 

Graphite         Combined 
Carbon             Carbon 

Silica 

Silver  gray 

3.13                     0.02 

5-5 

No.  2  soft 

3.48                    0.03 

3-5 

No.  i  soft 

3-53                0.03 

3-5*04.0 

No.   i   foundry 

3.49                0.07 

2.8103.5 

No.  2  foundry 

3-55               0.07 

2  .  2  tO  2  .  6 

No.  3  foundry 

3.48               o.io 

2  .  O  tO  2  .  4 

Gray  forge 

3.00                0.57 

1.3*01-7 

Alottled                                  2  .  1  1                 i  .  22 

i  .  i  to  i  .  6 

White                                    o.  10                 2.92 

O.7  tO  1.2 

The  table  shows  the  increase  of  combined  carbon  and  the  de- 
crease of  silicon  as  the  grade  progresses  toward  white  iron. 


OF  THE  COMMON  METALS.  233 

61.     IRON  BLAST-FURNACE  SLAGS. 

The  gangue  contained  in  the  ore,  as  well  as  the  ash  of  the  fuel 
added  to  the  blast-furnace  charge,  must  be  fluxed  by  the  addi- 
tion of  a  cheap  base,  such  as  lime  or  magnesia,  to  produce  a  fluid 
slag.  Now  the  ores  used  in  the  large  iron  districts  contain  little 
sulphur,  but  coke  may  carry  from  0.25%,  which  is  low,  to  2%, 
which  is  high,  or  on  an  average,  i%.  This  sulphur  cannot  be 
permitted  to  enter  the  iron,  while  by  using  a  slag  containing 
plenty  of  lime  and  a  hot  furnace,  it  may  be  forced  into  the  slag. 
Thus,  on  Lake  Superior  iron  ore,  and  using  coke,  a  suitable  slag 
had  the  composition  35.1%  SiO2,  14.2%  A12O3,  22.5%  CaO, 
22.4%  MgO,  while  the  iron  produced  contained  1.4%  Si  and 
0.05%  S.  In  another  instance,  in  a  fairly  hot  furnace,  a  slag  of 
36.1%  SiO2,  12.9%  A12O3,  41.7%  CaO,  and  7.3%  MgO,  took  up, 
principally  from  the  coke,  1.6%  S,  and  produced  an  iron  contain- 
ing 2.15%  Si  and  only  0.02%  S.  The  slag  was  quite  clean,  re- 
taining but  0.54%  Fe. 


PART  VI.    COPPER 


PART  VI.     COPPER. 

62.     ORES  OF  COPPER. 

In  describing  the  ores  of  copper,  it  must  not  be  for- 
gotten that  we  arc  to  think  of  them  as  mineral  aggregates 
(mixed  ores),  containing  often  less  than  10  to  15%  Cu, 
and  consequently,  carrying  a  proportion  of  associated  min- 
eral in  large  excess.  In  treating  such  ores  the  question  of  dispos- 
ing of  the  gangue  is  of  as  much  importance  as  obtaining  the  cop- 
per. Though  there  are  many  kinds  of  copper  ores,  those  of  com- 
mercial importance  are  few  in  number.  We  may  divide  them  into 
three  classes:  the  sulphides  (and  especially  chalcopyrite),  being 
the  most  common ;  the  oxides,  including  the  carbonates,  and  fin- 
ally rocks  containing  native  copper. 
Copper  sulphides. 

Chalcopyrite. — Cu  Fe  S2  contains  34.4%  Cu,  30.5%  Fe,  and 
35.1%  S.  This  is  by  far  the  most  widely  distributed  and  the  most 
abundant  of  the  ores  of  copper  and  furnishes  the  world's  principal 
supply  of  the  metal.  It  is  frequently  accompanied  by  an  excess 
of  iron  pyrite  with,  even  in  the  massive  sulphide,  a  certain  amount 
of  silicious  gangue.  In  consequence,  such  ores  ordinarily  carry 
but  a  small  amount  of  copper,  and  are  particularly  suited  to  pyrite 
or  semi-pyrite  smelting.  Silver  and  gold  are  also  to  be  found 
in  these  ores  in  small  quantity.  Thus,  at  the  deposits  of  Mt. 
Lyell,  Tasmania,  which  have  been  so  successfully  worked  by 
pyritic  smelting,  the  ore,  chiefly  a  massive  iron  pyrite  containing 
some  chalcopyrite,  carries  4.5  to  5%  Cu;  15  oz.  Au  and  3  oz. 
Ag  per  ton. 

Chalcocite  (Copper  glance). — Cu2S  contains  79.7%  Cu  and 
20.3%  S  when  pure.  Even  the  crystals  are  seldom  pure,  the 
copper  having  been  replaced  by  iron  and  other  metals,  but  still 
showing  the  characteristics  of  the  mineral  when  carrying  as  little 
as  55%  Cu.  When  pure,  however,  it  resembles  the  artificial 
product  of  the  furnace,  'white  metal,'  a  high-grade  matte. 


238  THE    METALLURGY 

Bornite  (Peacock  ore). — 3  Cu2S,  Fe2S3  containing  55.6%  Cu, 
16.3%  Fe,  and  28.1%  S.  Like  copper  glance,  this  mineral  is  far 
from  uniform  in  composition,  varying  in  richness  from  42  to  70% 
Cu,  without  losing  its  characteristic  varied  colors. 

Tetrahedritc  (Gray  copper,  fahlerz). — Cu2S,  FeS,  ZnS,  AgS, 
PbS  (Sb  S3As2S3)  containng  30.4%  Cu.  This  ore,  because  of 
its  obnoxious  contents,  antimony  and  arsenic,  is  an  unfavorable 
ore  of  copper ;  and  it  is  only  its  silver  contents,  when  high,  which 
justify  its  being  smelted  at  all. 

Copper  oxides  and  carbonates  (Oxidized  ores  of  copper). — 
These  ores  are  the  products  of  decomposition  of  sulphides  by  the 
elements.  As  a  result  we  find  such  ores  at  or  near  the  surface, 
accompanied  by  iron  in  oxidized  form  (gossan),  which  changes 
in  depth  to  the  sulphide,  both  of  iron  and  copper.  Being  in  oxi- 
dized form,  they  are  often,  when  of  low  grade,  suitable  ores  for 
the  extraction .  of  their  values  by  leaching  methods. 

Cuprite  (Red  oxide  of  copper). — Cu,O  containing  88.8%  Cu, 
11.2%  O.  This  mineral,  a  product  of  decomposition,  often  per- 
meates large  masses  of  iron  oxide.  Large  lumps  of  it  sometimes 
occur  which  are  evidently  the  result  of  oxidization  of  native  cop- 
per, the  centre  still  remaining  unaltered. 

Melaconitc. —  (Black  oxide  of  copper)  CuO,  containing 
79.8%  Cu2,  20.2%  O.  This  ore,  with  its  copper  replaced  in  part 
by  oxides  of  iron  and  manganese,  is  sometimes  found  in  masses 
large  enough  to  pay  for  extraction,  and  containing  20  to  50%  of 
the  metal.  The  so-called  black  oxide  of  the  Blue  Ridge  region, 
on  the  border  between  Tennessee,  North  Carolina  and  Virginia, 
seems  to  be  an  intimate  mixture  of  copper  glance,  copper  oxide, 
copper  carbonate,  native  copper  and  iron  oxide  and  sulphide. 
Such  ore  is  readily  pile-roasted. 

Malachite. — CuCO3,  Cu(H2O)  will  carry,  when  pure,  57.3% 
Cu.  This  ore  occurs  widely  distributed,  ordinarily  in  non-paying 
quantities  as  a  decomposition  product  in  surface  deposits,  but 
sometimes  rich  enough  for  a  workable  ore.  It  is  found  mixed  with 
lime,  dolomite,  oxides  of  iron  and  manganese,  and  silica  in  its 
various  forms.  It  is  very  difficult  to  judge  from  external  appear- 
ance of  the  content  of  the  ore  in  copper. 


OF    THE    COMMON    METALS. 


239 


FIG.  104.    BLAST-FURNACE  FOR  OXIDIZED  COPPER  ORES. 


24O  THE    METALLURGY 

Azuritc. — 2  Cu  CO3  -j-  Cu  2  H2O  has  55.2%  Cu  when  pure. 
This  striking  looking  ore  occurs  much  the  same  as  malachite,  but 
less  abundantly,  coloring  other  oxides. 

Xatii'c  copper. — Cu. — It  is  found  very  extensively  in  the  copper 
region  of  Lake  Superior.  Elsewhere  it  is  found  too  sparingly  to 
have  commercial  value.  In  the  Lake  Superior  country  it  is  found 
in  wide  lodes  or  veins  disseminated  through  the  rock,  in  such  a 
way  that  it  forms  from  only  0.5  to  4%  of  the  whole  mass  and 
consequently  all  the  vein  matter  must  be  concentrated  to  obtain 
the  copper.  The  concentrate,  or  'mineral,'  as  it  is  locally  named, 
will  carry  from  30  to  95%  copper,  the  lower  grades  still  holding 
considerable  gangue. 

63.     SMELTING  THE  OXIDIZED  ORES  OF  COPPER. 

In  the  southwest  of  the  United  States  (New  Mexico  and  Ari- 
zona) are  to  be  found  many  oxidized  ores,  containing  copper  as 
cuprite  and  malachite,  accompanied  sometimes  by  small  amounts  of 
sulphides,  as  copper  glance.  These  ores  have  been  treated  for 
the  winning  of  their  copper  in  one  operation  in  blast-furnaces,  by 
a  reduction  smelting  for  metallic  copper.  The  same  method  is 
followed  in  the  Lake  Superior  copper  country  for  the  smelting 
of  copper-bearing  reverberatory  slags  made  in  the  smelting  of 
the  native  copper  'mineral'  of  that  country.  These  slags  contain 
some  copper  as  copper  oxide,  together  with  shot  and  scrap  copper, 
removed  during  the  skimming. 

Fig.  104  and  105  represent  an  exterior  and  a  vertical  sectional 
elevation  of  the  blast-furnace  or  cupola  used  for  these  operations. 

The  water-jacket  (often  made  in  sections)  is  of  flange  steel. 
It  is  surrounded  by  a  'wind  box'  connected  to  the  blower  by  means 
of  a  blast  pipe,  shown  in  the  exterior  view.  The  tuyeres  are 
within  the  water  space  of  the  jacket,  and  have,  opposite  them, 
mica-covered  peep-holes  for  observation  of  their  condition.  The 
curb,  extending  from  the  base  plate  to  the  jackets,  forms  the 
sheet-iron  covering  of  the  crucible.  The  base  plate  rests  upon 
four  iron  columns,  which  are  supported  by  a  solid  foundation. 
The  base  plate,  together  with  two  hinged  drop-doors,  forms  the 


OF    THE    COMMON    METALS. 


24I 


FIG.  105.     BLAST-FURNACE  FOR  OXIDIZED  COPPER  ORES   (SECTION). 


242  THE    METALLURGY 

bottom  of  the  crucible.  The  crucible  is  a  receptacle,  lined  on  the 
sides  with  fire-brick,  and  on  the  bottom  with  a  mixture  called 
'steep/  made  of  coke  dust,  ground  fire-brick  and  clay,  or  with  fire- 
brick only.  It  forms  an  infusible  lining  to  protect  the  iron  ex- 
terior, and  to  hold  or  retain  the  heat  of  its  contents.  It  is  this 
crucible  which  holds  the  molten  portion  of  the  charge,  the  slag 
separating  out  and  floating  above  the  molten  copper,  which  has 
been  reduced  from  the  ore.  There  are  two  tap  holes  to  the  cru- 
cible, the  lower,  close  to  its  bottom,  for  the  removal  of  the  molten 
'blister'  copper,  and  the  upper  one  for  removal  of  the  supernatant 
slag.  From  time  to  time,  as  either  of  these  products  accumu- 
lates in  the  crucible,  it  is  tapped  out  by  punching  a  hole  through 
the  clay  stopping  of  the  tap-hole  with  a  pointed  tapping-bar.  The 
flow  is  stopped  again  by  means  of  a  plug  of  clay  stuck  on  the  end 
of  a  button-ended  stopper-rod  or  dolly.  The  slag  is  received  into 
a  slag  pot,  shown  in  Fig.  106,  and  pushed  out  to  the  dump  by  a 
man  who  pours  out  its  contents  over  the  edge  of  the  dump. 

The  blister  copper  is  received  in  a  copper-bullion  pot  or  mold, 
like  the  slag  pot,  mounted  on  wheels.  As  shown  (Fig.  107)  the 
water-jacket  forms  the  wall  of  the  shaft  of  the  furnace,  and  within 
this  space,  the  fusion  of  the  charge  takes  place.  Formerly,  this 
wall  was  of  fire-brick,  but  it  was  found  that  it  was  soon  eaten 
through,  or  corroded,  by  the  action  of  the  molten  slag  upon  it. 
The  water-jacket  shown  as  a  double  annular  shell,  within  which 
water  freely  circulates,  prevents  such  corrosion,  permitting  un- 
interrupted operation.  Cold  water  enters  the  jacket  in  a  regu- 
lated stream  to  replace  heated  water  flowing  out  at  a  temperature 
of  70  to  80°  C.  It  will  be  noticed  that  the  shaft  of  the  furnace 
widens  out  as  it  goes  upward,  forming  a  bosh.  Above  the  shaft 
rises  the  stack  or  hood  to  carry  off  the  fumes  from  the  charge. 
A  feed  door,  shown  in  the  hood,  is  provided  to  introduce  the 
charge.  It  is  to  be  understood,  that  when  in  operation,  the  fur- 
nace is  kept  full  to  the  feed  door  with  charge  and  fuel.  These  are 
placed  in  layers,  first  a  layer  of  fuel  and  upon  it  a  layer  of  charge, 
and  so  alternately.  The  blast  rises  through  this  mass  of  materials 
called  'smelting  column,'  under  a  pressure  of  12  oz.  to  the  square 
inch.  The  entering  air  causes  an  intense  combustion  of  the  coke 
and  fuses  the  charge.  In  contact  with  the  glowing  coke,  reduc- 


OF   THE    COMMON    METALS. 


243 


244  THE    METALLURGY 

tion  is  also  effected,  by  which  the  copper  separates  out  in  drops, 
and  finds  its  way  to  the  bottom  of  the  crucible.  In  the  oxidized 
ores  of  copper  may  often  be  found  copper  glance  or  other  sulphide. 
This  sulphur,  quite  small  in  amount,  is  often  dissipated  by  the 
blast,  but  otherwise,  uniting  itself  to  some  of  the  copper,  it  may 
make  a  small  amount  of  matte  of  about  70%  copper.  This  matte 
is  to  be  found  as  a  thin  coating  or  shell  upon  the  upper  surface 
of  the  blister  copper  when  it  cools  in  the  copper-bullion  mold. 

The  separation  between  slag  and  copper  in  the  crucible  is  not 
perfect,  but  drops  of  copper  are  liable  to  escape  with  the  slag. 
To  save  this  a  forehearth  or  brick-lined  iron  box  is  placed  to 
catch  the  flow  from  the  slag  spout.  The  slag  flows  into  the  fore- 
hearth  (which  is  full  of  molten  slag)  on  one  side,  leaving  by  a 
spout  on  the  opposite  one.  Any  drops  of  copper,  in  the  gradual 
passage  of  the  slag  across  the  forehearth,  settle  out  to  the  bottom. 
When  the  copper  has  accumulated  in  sufficient  quantity  in  the 
forehearth,  it  may  be  tapped  off  by  a  lower  spout  shown  on  the 
side. 

As  a  result  of  the  exterior  cooling  influences,  the  crust  of  slag, 
which  forms  upon  the  borders  of  the  forehearth,  becomes  thicker, 
and,  after  several  days,  may  result  in  the  central  molten  pool  of 
slag  becoming  too  small  for  effecting  a  good  separation  of  copper 
from  slag.  When  this  happens,  the  forehearth  may  be  pried  out 
of  the  way  on  its  wheels  and  another  one  run  into  position.  For  a 
small  round  blister  copper  furnace,  such  as  already  illustrated,  a 
settling  pot  may  be  used  in  place  of  a  forehearth.  As  shown  in  Fig. 
109,  it  resembles  a  large  slag  pot,  but  has  spouts  for  the  removal  of 
the  slag,  and  when  copper  accumulates  in  the  bottom  this  settling 
pot  may  be  readily  replaced  by  another.  The  copper  in  the  former 
settler  can  be  removed  when  the  contents  have  stood  and 
solidified. 

The  cost  of  treatment  of  these  ores  in  the  Southwestern  United 
States  may  be  given  at  $8  per  ton.  The  slag  generally  runs  from 
1.5  to  2.5%  in  copper,  so  that,  under  the  best  conditions,  we  may 
rely  on  a  recovery  of  but  82 9t  of  the  copper,  while  the  extraction 
of  the  silver  in  the  ore  will  go  to  91  % .  This  considerable  loss  of 
the  copper  in  the  slag  is  one  of  the  drawbacks  of  the  process.  It 
would  appear  that,  as  has  already  been  demonstrated,  the  use  of 


OF    THE    COMMON    METALS. 


245 


246 


THE    METALLURGY 


hard  coal  would  give  much  better  reduction — to  as  low  as  0.6  to 
i%  Cu  as  in  the  Lake  Superior  copper  country. 

64.     STORAGE  OF  MATERIALS  OF  THE  CHARGE. 

Sometimes  the  arrangements  for  storing  the  supplies  are  very 
simple,  they  being  put  in  heaps  or  in  simple  bins.    However,  it  is 


FlG.    108.        FOREHEARTH. 

becoming  customary  to  put  the  most  used  of  such  material  in  hop- 
per-bottomed storage  bins  whence  they  may  be  drawn  off  to  the 
charge-cars  or  buggies  by  means  of  gates  and  chutes.  The  disad- 
vantage of  the  method  is  that  the  ore.  if  lumpy,  varies  greatly,  be- 
ing sometimes  coarse  and  sometimes  fine,  and  making  correspond- 
ing variation  in  the  operations  of  the  furnace.  Coke  is  generally 


OF    THE    COMMON    METALS. 


247 


248  THE    METALLURGY 

stored  in  heaps.  There  should  be  a  good  supply,  both  of  it  and  of 
the  fluxes  kept  in  storage,  so  that  in  case  of  short  supply  or  failure 
in  delivery,  the  storage  can  be  drawn  on.  Such  a  reserve  stock 
need  not  cost  much,  and  will  well  repay  the  outlay  in  the  in- 
creased certainty  of  running. 

While  in  copper  smelting  it  is  possible  to  estimate  the  needed 
ingredients  by  the  shovelful,  it  is  better  to  weigh  everything 
which  goes  to  the  furnace,  except  the  foul  slag  or  old  charge  re- 
turned in  the  regular  running.  The  latter  has  been  charged  once 
to  the  furnace,  and  should  not  be  again  reckoned  in.  The  number 
of  charges  is  recorded  by  keeping  tally  of  them  on  a  board  pierced 
with  100  holes,  set  in  ten  rows.  The  weighing  of  the  charge  is 
done  by  the  charge-wheelers,  and  the  metallurgist  should  con- 
stantly see  that  weights  and  charges  are  correctly  and  properly 
registered.  These  data  form  an  important  feature  of  the  question 
of  costs.  The  roasted  ore,  when  made  at  reverberatory  roasting 
furnaces,  is  dumped  out  on  a  cooling  floor  and  from  it  is  shoveled 
directly  into  the  charge-buggies.  The  fuel  is  put  on  the  charge- 
plate  at  one  side  of  the  furnace  and  the  charge  on  the  other,  and 
this  arrangement  is  alternated,  the  following  charge  being 
dumped  where  the  coke  had  just  lain.  It  is  to  be  understood  that 
the  fuel  and  charge  are  delivered  into  the  furnace  in  alternate 
layers.  The  coke  is  spread  evenly  over  the  surface  of  the  ma- 
terials in  the  furnace  and  the  charge  itself  fed  in  so  as  to  get 
lumps  to  the  centre  and  fine  to  the  sides,  taking  care,  however,  to 
feed  so  that  the  blast  conies  through  evenly.  The  slag  from  an 
exhausted  forehearth — that  which  runs  out  with  matte,  or  from 
cleaning  out  a  spout,  etc.,  all  contains  drops  of  matte,  and  is 
called  foul  slag.  It  is  to  be  returned  to  the  furnace  as  a  welcome 
addition  to  the  charge. 

65.     RELATIVE  ADVANTAGES  OF  BLAST-FURNACE  AND 
REVERBERATOR v  SMELTING. 

The  output  of  the  blast-furnace  is  larger  than  that  of  the  rever- 
beratory. 

The  blast-furnace  slag  is  generally  cleaner.  The  blast-furnace 
takes  its  charges  continuously,  and  the  removal  of  slag  is  practi- 
cally continuous.  Matte  is  removed  periodically. 


OF    THE    COMMON    METALS.  249 

The  blast-furnace  takes  less  fuel  than  the  reverberatory,  though 
that  fuel  is  more  costly  per  ton  than  reverberatory  coal.  In 
pyritic-smelting  but  little  coke  is  used. 

The  reverberatory  can  smelt  finely  ground  ore  which  would 
create  much  flue-dust  in  the  blast-furnace. 

Where  coal  is  cheap  the  fuel  cost  may  be  lower  in  reverberatory 
than  in  regular  blast-furnace  work. 

Slags  in  the  reverberatory  may  be  more  infusible,  in  fact  such 
slags  may  contain  pieces  of  unfused  silicious  ore,  which  never- 
theless have  been  freed  from  gold  and  silver  values. 

Repairs  to  the,  reverberatory  are  greater  than  in  blast-furnaces. 
Both  kinds  of  furnaces  have  their  troubles,  the  blast-furnace  in 
freezing  up,  the  reverberatory,  through  difficulty  in  reaching 
fusion  temperature. 

The  slag  and  matte  are  removed  periodically  from  the  rever- 
beratory, which  occasions  delay  and  consequent  cooling  of  the 
furnace. 

The  grade  of  matte  is  increased  in  the  reverberatory  owing  to 
the  removal  of  sulphur  by  the  following  reactions   which  take 
place  when  the  charge  is  fused  down. 
FeS  +  3  Fe2O3  +  7  SiCX  =  7  FeO  SiCX  +  SCX 
Ba  SO4  +  SiCX  =  BaO  SiO,  +  SO3 

That  is,  where  ferric  iron  or  barium  sulphate  (heavy  spar)  is 
present,  sulphur  is  eliminated. 

66.     CALCULATION  OF  CHARGE  IN  REGULAR  COPPER  BLAST- 
FURNACE MATTING. 

A  charge,  suited  to  such  smelting,  would  consist  of  roasted  ore 
with,  or  without,  the  addition  of  oxidized  copper  ores,  and  of 
silicious  ores,  in  case  it  should  be  desirable  to  obtain  the  gold  and 
silver  from  them.  The  products  of  the  furnace  are  slag  and 
matte,  the  former  being  the  result  of  the  union  of  the  silica  in  the 
charge  with  various  bases,  namely  iron  and  lime ;  the  latter  be- 
ing that  complex  artificial  sulphide  due  to  the  sulphur  in  the 
charge,  taking  up,  first  copper,  and  then,  if  it  needs  it,  a  portion  of 
the  iron  present.  The  remainder  of  the  iron  enters  the  slag.  A 
copper  furnace  slag  may  be  quite  variable  in  composition,  the 


250  THE    METALLURGY 

main  requirement  being  that  it  shall  be  fusible.  Slags  varying 
from  25  to  40%  in  silica  are  found  in  copper  practice,  the  latter 
limit  being  somewhat  exceeded  where  bases  for  fluxing  add  much 
to  the  cost  of  smelting.  In  lead-silver  smelting  this  is  not  the 
case,  and  it  is  necessary  to  proportion  the  charge  so  as  to  have 
slags  of  a  type,  such  that  they  shall  be  lead-and-silver-free. 

As  a  simple  case,  let  us  take,  for  example,  1,000  Ib.  of  a  roasted 
ore  having  the  composition  5%  Cu,  25%  SiO2,  30%  Fe  and 
roasted  to  leave  in  it  10%  S.  This  is  to  be  smelted  with  limestone 
to  produce  a  slag  to  contain  35%  SiCX,  the  bases  FeO  and  CaO 
amounting  together  to  55%,  which  will  leave  10%  for  all  other 
elements.  The  limestone  contains  4%  SiO2  and  52%  CaO,  while 
the  coke  has  12%  ash,  this  ash  having  60%  SiO2,  10%  Fe  and 
15%  CaO.  These  figures  in  the  ash  correspond  to  7.2%  SiO2, 
1.2%  Fe  and  i.S%  CaO  in  the  coke.  A  metallurgist,  accustomed 
to  running  certain  ores,  knows,  with  some  degree  of  approxima- 
tion, how  much  flux  he  needs,  but  even  if  not,  he  makes  a  guess 
as  to  the  quantity  needed,  and  puts  the  figure  down  in  the  column 
of  dry  weights  of  the  charge  sheet  on  the  line  for  limestone. 
Say  this  is  300  Ib.,  and  that  he  has  decided  upon  10%  coke 
as  the  fuel  needed  for  this  charge  of  1,300  Ib.  or  130  Ib., 
which  is  also  set  down.  The  percentages  are  then  written 
in  their  proper  columns,  and  the  multiplications  for  the 
weights  in  pounds  carried  out,  neglecting  fractions.  At  the 
left  of  the  charge-sheet  write  in  the  slag  composition.  Obtain 
now  the  ratio  of  the  bases  to  silica,  as  shown,  giving  a  factor  of 
1.57  in  this  particular  case.  On  the  right  of  the  charge-sheet 
write  the  matte  composition  which,  in  the  case  of  mattes 
such  as  are  produced  in  blast-furnaces,  is  of  the  average 
composition  there  given.  Also  work  out  the  ratio  of  bases 
to  sulphur  giving  the  factor  3.  Let  us  take  up  the  sulphur. 
On  well-roasted  ore  we  may  reckon  on  a  loss,  by  vola- 
tilization, of  25%,  leaving  75%  to  go  to  the  formation  of 
matte.  The  75  Ib.  found,  multiplied  by  the  factor  3,  gives  225 
Ib.  of  copper  and  iron  needed  for  the  matte.  There  is,  however, 
4  Ib.  copper  lost  in  the  slag,  to  be  deducted  from  the  100  Ib.  given 
on  the  charge-sheet.  The  total  slag  equals  the  271  Ib.  SiO2  of 
the  charge-sheet,  divided  by  35%  SiO2  or  770  Ib.  Allowing 


OF    THE    COMMON    METALS. 


251 


0.5%  Cu  in  the  slag,  this  makes  4  Ib.  copper  so  lost,  leaving  96 
Ib.  copper  for  the  matte.  Then  225  —  96=129  Ib. 
of  Fe  to  be  taken  from  the  total  of  302,  leaving  173  Ib.  to  enter 
the  slag.  But  the  iron  in  the  slag  exists  as  FeO  or  as  56  Fe  to  72 
FeO,  or  as  7  to  9,  giving  us  222  Ib.  FeO.  To  this  add  the  CaO 
( 158  Ib.),  making  380  Ib.  of  both  bases.  Multiplying  the  silica,  271 
Ib.,  by  the  factor  1.57,  we  find  that  we  need  425  Ib.  of  the  bases 
FeO  and  CaO,  so  that  we  have  45  Ib.  too  little.  Now,  since  lime- 
stone is  approximately  one-half  CaO,  this  will  mean  the  addition 
of  about  90  Ib.  to  the  charge,  making  390  Ib.  as  the  proper  figure. 
Erase  the  original  figure  (300)  and  replace  with  the  amended 
amount  (390),  erase  where  needed,  and  recalculate  the  charge 
throughout.  This  time  we  will  come  to  within  a  few  pounds  of 
the  correct  amount.  The  amount  may  be  again  amended  without 
recalculating,  or,  when  within  5  or  10  Ib.  of  the  exact  figure,  may 
be  accepted  as  being  close  enough.  Variations  in  the  ores  in  the 
weighing  and  in  the  sulphur  volatilized  may  far  exceed  such 
minor  differences; 

Charge  Sheet. 


Weight 

Cu 

SiO2 

Fe&Mn 

CaO 

&MgO 

S. 

H2O 

Wet 

Dry 

t 

Wt. 

* 

Wt. 

% 

Wt. 

# 

Wt. 

< 

Wt. 

Roasted  Ore 

1000 

5.0 

100 

25.0 

250 

30.0 

300 

10.0 

100 

Limestone 

300 

4.0 

12 

52.0 

156 

Coke 

130 

7.2 

9 

1.2 

2 

1.8 

2 

100 

271 

302 

158 

100 

Cu  in  Slag     = 

4 

For  Matte  = 

129 

75 

Cu  in  Matte  = 

96       For  Slag     = 

173 

S  in 

Matte  = 

75 

225 

3 

Fe  in  Matte  = 

129                           Cu  and  Fe  in  Matte  =  '  225 

SLAG 

SiO2  = 

FeO  &  CaO   =  55.0$ 

FeO  &  CaO        .  „  ,,  , 
— =  1.57  (factor) 


173  Fe  =  222  =  FeO 
158  =  CaO 

Actual  FeO  &  CaO  =  380 
Needed    "  "     —425 

CaO  too  little          =    45 


MATTE 

S  =  23.0 
Cu  &  Fe  =  69.0 


Dividing  the  225  Ib.  Cu  and  Fe  by  their  percentage  69,  we  get  a 
factor  3.3.     The  pounds  of  Cu  and  Fe,  divided  by  this  factor, 


252  THE    METALLURGY 

give  Cu  29%  and  Fe  40%  as  the  proportions  of  those  metals  in 
the  matte.  In  a  similar  manner  we  may  find  the  amount  of  FeO 
and  CaO  in  the  slag. 

The  metallurgist  can  seldom  count  upon  the  charge  coming 
down  as  calculated ;  and  as  soon  as  it  is  down  he  should  take  a 
sample  of  the  slag  for  analysis.  It  should  be  determined  for  Cu, 
SiO2,  Fe  and  CaO.  These  determinations  will  take  upward  of 
two  hours,  and  from  them  the  furnace  charge  can  be  corrected 
as  desired,  or,  if  the  slag  appears  satisfactory,  and  is  low  in  cop- 
per, it  may  be  left  unchanged. 

67.     MATTE  AND  SLAG. 

Matte  is  a  complex  artificial  sulphide  formed  in  smelting  opera- 
tions as  the  result  of  the  union  of  sulphur  with  bases.  Iron  sul- 
phide (FeS),  such  as  is  used  in  the  generation  of  H2S  in  the 
laboratory  is  its  simplest  form.  When  part  of  the  iron  is  substi- 
tuted by  copper,  copper  matte  Cu,FeS  is  formed.  If  lead  enters 
this  compound,  as  is  the  case  in  lead  smelting,  we  will  have 
a  leady  copper-matte  Cu,Pb,FeS.  In  smelting  copper-nickel  ores 
a  matte  Xi,Cu,FeS  is  produced.  It  must  be  understood,  how- 
ever, that  these  combinations  are  not  in  relation  to  the  atomic 
weights.  Thus,  in  copper  mattes  we  may  find  the  copper  in 
any  proportion  from  little  up  to  75%.  Normal  mattes  carry 
20  to  26%  sulphur,  the  remainder  being  made  up  of  the  metals. 
It  will  be  noticed  that  where  there  is  copper  present  in  a  smelting 
charge,  it  will  be  taken  up  by  the  sulphur  first;  and,  then  if  the 
sulphur  needs  other  base,  it  will  take  up  iron.  In  other  words  the 
affinity  of  sulphur  for  copper  is  greater  than  for  iron. 

As  compared  with  ore,  when  roasted,  matte  retains  its  sulphur 
more  firmly,  and  needs  longer  time  and  more  heat  in  roasting. 
Thus,  a  matte,  when  well  roasted,  may  still  hold  5  to  6%  S, 
while  the  correspondingly  roasted  ore  would  retain  but  3  or  4% 
S.  Mattes  vary  in  specific  gravity.  A  matte  high  in  copper  is 
heavy  as  compared  with  a  normal  50%  copper  matte,  which  latter 
would  have  the  specific  gravity  of  5.  Mattes  often  take  up  zinc 
from  the  charge,  becoming  lighter  in  so  doing.  On  the  other 
hand  we  may  have  the  case  of  slags  high  in  iron  of  a  specific 


OF    THE    COMMON    METALS.  253 

gravity  of  3,  which,  with  light  mattes,  result  in  a  poor  separation 
because  of  the  less  difference  in  specific  gravity.  On  the  other 
hand,  silicious  limy  slags  of  low  specific  gravity  promote  separa- 
tion. 

Copper  furnace  slags. — Very  considerable  variations  in  slag 
composition  are  permissible  in  copper-smelting  blast-furnace  prac- 
tice, the  requirement  being  that  the  slag  should  be  fusible,  and 
fluid  enough  to  flow  from  the  tap-hole  of  the  furnace.  Slags  hav- 
ing the  minimum  and  maximum  contents  in  silica  and  bases 
shown  herewith,  can  be  successfully  used  in  the  blast-furnace. 
Of  course,  when  they  are  at  a  minimum  in  one  element,  they  are 
at  a  maximum  in  others. 

SiO2       FeO     Al2Oo       CaO     MgO       ZnO          Ba 
Minimum  20  2  2  2  o  o  o 

Maximum          57  70  18  40  8  20  42 

68.     COMPOSITION  OF  COPPER  MATTE. 

In  the  following  is  given  the  composition  of  five  grades  of 
copper  matte. 

I     28%  Cu  35%  Fe  ,  23%  S  (coarse  metal) 
II     35%    "  30%  "     23%  " 

III  51%    "   1 8%   "     23%   "  (shipping,  or  converter  matte) 

IV  60%    "  13%   "     23%  "blue  metal 
V    70%    "     5%  "     20%     "      " 

It  will  be  noticed  that  in  the  higher-grade  mattes,  because  of 
the  small  amount  of  Fe  present,  the  heat  in  converting  is  not  so 
well  kept  up.  However,  the  Cu2S  becomes  in  part  oxidized  to 
Cu2O  and,  reacting  on  the  remaining  cuprous  sulphide,  gives 
copper,  and,  the  reaction  being  exothermic,  keeps  the  molten  con- 
tents of  the  converter  liquid. 

We  have 

1 i )  2  Cu2S  +  6  O  =  2  Cu2O  +  2  SO2 

2X20200  2X42000  2X71000=  +  185600 

(2)  Cu2S  +  2  Cu20  =  6  Cu  +  SO2 

20200         2X42000  2X71000= 23,600 

Or,  as  the  result  of  the  two  reactions,  185600 — 23600=162000,  or 
per  pound  of  copper  638  pound  calories. 


254 


THE    METALLURGY 


69.     THE  COPPER-MATTING  BLAST-FURNACE. 


For  producing  matte  from  copper-bearing  ores,  whether  roasted 
or  raw,  a  furnace  having  a  simple  hearth  but  no  crucible  is  used, 


FIG.  110.     COPPER  MATTING  FURNACE. 

as  shown  in  the  illustration,  Fig.  no,  a  being  a  transverse  and  b 
a  longitudinal  elevation  of  a  furnace  42  by  120  in.  hearth  dimen- 
sions. Fig.  in  is  a  perspective  view  of  the  lower  part  or  iron- 
work of  the  same  furnace.  The  sole-plate  of  the  hearth  is  sus- 


OF    THE    COMMON    METALS. 


255 


tained  by  jack-screws  which  stand  upon  the  foundations,  and, 
when  desired,  it  can  be  lowered  and  removed  altogether  for  re- 
pairs. The  sole-plate  is  protected  from  the  action  of  molten  ma- 
terials by  a  layer  of  fire-brick  8  in.  thick,  forming  an  inverted 
arch  abutting  on  the  side  water- jackets,  and  upon  it  stands  the 


FIG.  111.     PERSPECTIVE  VIEW  OF  COPPER  MATTING  FURNACE. 

lower  tier  of  water-jackets,  more  particularly  shown  in  Fig.  112. 
These  jackets,  9  ft.  high,  are  pierced  for  the  tuyeres,  have  inlet 
openings  at  half  the  height  of  the  bosh  for  water  supply,  and  out- 
lets at  their  highest  point,  so  that  they  may  be  kept  completely 
filled  with  water.  They  are  clamped,  or  tied  together  with  heavy 
angle-irons.  In  section  b  Fig.  113  is  shown  a  water-cooled  spout, 


256 


THE    METALLURGY 


OF    THE    COMMON    METALS. 


257 


by  which  the  slag  and  matte  together  flow  from  the  furnace.  The 
slag  before  overflowing  fills  the  spout  and  covers  its  outlet  from 
the  furnace,  thus  preventing  the  escape  of  the  blast;  and  as  fast 
as  it  forms  it  escapes  in  a  regular  stream.  On  the  side,  at  the 
middle  jacket,  is  to  be  seen  another  outlet  furnished  with  a  spout, 
generally  kept  closed,  but  which  is  opened  when  it  is  desired  to 
empty  the  hearth  of  its  contained  matte  and  slag.  Both  these 
apertures  are  closed,  except  for  the  tap-hole  openings,  by  a 


FIG.  113.     TRAPPED  SPOUT. 


water-cooled  tap-jacket.  The  lower  jackets  are  suspended  from 
I-beams,  so  that  the  hearth  may  be  removed  without  disturbing 
them.  Above  the  tuyeres  the  side- jackets  widen  out  forming 
a  I4~in.  bosh  or  enlargement  of  the  furnace,  so  that,  at,  their  top, 
they  are  5  ft.  6  in.  wide.  Above  the  lower  jackets  is  another  tier, 
above  which  comes  the  cast-iron  distributing-plates  forming  the 
sill  of  the  feed-doors.  It  will  be  noticed  that  the  feed-doors  ex- 
tend the  whole  length  of  the  furnace  and  on  both  sides,  thus  mak- 
ing the  whole  interior  accessible,  not  only  to  the  feeding  and 
trimming  of  the  charges,  but  also  for  cutting  off  whatever  accre- 
tions may  form  upon  the  interior  surfaces  of  the  jackets.  The 


258  THE    METALLURGY 

upper  portion,  being  the  stack  or  closed  top  of  the  furnace,  is  of 
brick,  carried  upon  a  deck  or  mantel-plate  of  I-beams,  these  again 
resting  on  the  columns  which  are  firmly  fixed  in  the  foundation. 
The  upper  portion  of  the  stack  is  of  sheet-steel,  and  is  completed 
by  a  circular  pipe  extending  through  the  roof  of  the  furnace 
building.  The  bustle-pipe,  by  which  the  blast  is  brought  to  the 
tuyeres,  extends  around  three  sides  of  the  furnace,  and  connects 
by  sheet-metal  blow-pipes  to  the  tuyeres  shown  more  in  detail  in 
Fig.  114.  In  this  figure  will  be  noticed  the  method  of  holding 
the  tuyere  against  the  jacket  by  a  tie-bolt  on  each  side.  The  front 
of  the  tuyere  has  a  cap,  \vhich  may  be  removed  for  access  to  the 
interior,  and  a  mica-covered  peep-hole  through  which  the  condi- 
tion of  the  furnace  can  be  observed.  Each  tuyere  has  its  own 
shut-off  valve. 

Just  below  the  bustle-pipe  will  be  seen  a  waste-launder  or 
trough,  which  receives  the  overflow  from  the  jackets,  and  below 
that  again  comes  a  3~in.  water-supply  pipe  branching  to  each 
jacket,  both  lower  and  upper,  and  to  the  water-cooled  spout  at  the 
front,  more  particularly  shown  in  Fig.  113. 

70.     BLAST-FURNACE  MATTE  SMELTING  OF  COPPER. 

The  smelting  of  copper-bearing  ores  for  the  production  of 
matte  may  be  divided  into  regular  smelting  and  pyrite  smelting. 
The  former  method  is  used  when  the  quantity  of  sulphur  present 
in  the  charge  is  small,  as  for  instance  when  using  partly  oxidized, 
roasted  or  silicious  ores.  Pyrite  smelting  applies  with  a  charge 
having  a  high  content  in  sulphur,  and,  in  which  the  sulphur  is 
incidentally  burned,  furnishing  heat  in  so  doing. 

71.     REGULAR  MATTE  SMELTING. 

This  method  has  an  advantage  in  that  its  operation  is  much 
more  certain  and  regular  than  is  the  case  with  pyrite  smelting. 
Ores,  carrying  much  sulphur  must,  however,  be  roasted,  and  since 
blast-furnace  smelting  is  not  suited  to  fine  ores,  either  heap  or 
stall-roasted  ores  should  be  used.  With  these  ores  there  can  also 
be  used  silicious  and  oxidized  ores.  The  amount  of  matte  made 
(matte  fall)  depends  upon  the  percentage  of  sulphur  in  the 


OF    THE    COMMON    METALS.  259 

charge,  so  that  to  get  a  good  concentration  (ratio  of  ore  to  matte 
produced)  the  sulphur  should  be  kept  low.  To  the  ores  just  speci- 
fied should  be  added  limestone  or  iron  ore  for  flux,  together  with 
10  to  15%  of  fuel,  generally  coke. 

Starting  and  operation  of  the  blast-furnace.  Since  a  modern 
blast-furnace  is  water-jacketed,  the  operation  of  warming  it  up  is 
a  simple  one.  The  crucible  being  lined  with  fire-brick  can  be  soon 
dried  out  and  warmed  by  several  hours'  heating  with  a  wood  fire. 
The  tap-openings  are  left  open  to  permit  entrance  of  the  air  to 


FIG.  114.     TUYERE. 

the  fuel.  The  hearth,  being  hot,  the  ashes  are  scraped  out  and  a 
fresh  fire  of  wood  is  put  in,  filling  the  interior  about  half  way  up 
to  the  tuyeres  (12  in.  deep).  Be  careful  to  have  the  wood  uni- 
form in  size,  so  that  it  may  burn  away  uniformly.  Upon  the 
wood  is  scattered  some  charcoal,  if  you  have  it,  to  just  cover  the 
wood.  Upon  this  is  put  a  layer  of  coke  coming  to,  say  il/2  to  2 
feet  above  the  tuyeres.  Care  must  be  taken  to  get  this  bed  of  fuel 
burning  uniformly  by  checking  the  draft  at  the  front  and  promot- 
ing it  at  the  rear,  if  necessary.  We  can  now  begin  charging. 
Supposing  that  the  charge  is  to  be  2,000  lb.,  and  that  we  are  to  use 
12%  of  coke,  or  240  lb.  per  charge.  We  put  in  the  240  lb.,  and 
upon  it  500  lb.  fusible  slag,  then  several  more  charges  each  con- 
sisting of  240  lb.  coke  with  1,000  lb.  slag.  Following  we  should 
put  in  a  number  of  charges  more,  each  of  240  lb.  fuel  and  2,000 


260  THE    METALLURGY 

lb.  slag.  The  charges  of  slag  should  be  of  weight  sufficient  en- 
tirely and  promptly  to  fill  the  forehearth.  We  may  now  begin 
putting  in  the  regular  charges  which  have  already  been  calculated. 
At  this  time,  the  fuel  being  fully  ignited,  the  blast  may  be  turned 
on,  gently  at  first,  and  gradually  increasing  for  half  an  hour,  when 
the  furnace  should  be  in  full  blast.  An  extra  force  of  men  should 
be  busy  charging,  so  as  to  get  the  furnace  filled  as  rapidly  as  pos- 
sible. At  the  slag  floor  before  the  blast  has  been  turned  on,  the 
slag-spout  should  have  been  put  up  and  the  tap-opening  closed 
with  balls  of  the  plugging  material.  This  stopper-clay  is  some- 
times to  be  obtained  from  a  neighboring  clay-bank,  and  is  some- 
times made  of  a  mixture  of  coarsely  ground  fine  brick  and  clay. 
The  brick-lined  forehearth  has  been  warmed  up  at  the  same  time 
as  the  furnace.  It  would  be  well  to  have  openings  low  down  on 
the  sides  to  permit  the  entrance  of  air,  which  may  be  plugged  at 
the  last  moment.  Thus  the  fire  does  not  get  sluggish.  Again,  the 
fuel  on  the  forehearth  should  be  set  round  against  the  walls,  and 
not  at  the  middle  of  the  hearth.  Use  the  wood  moderately,  and  as 
it  burns  away  and  as  charcoal  and  ashes  accumulate,  shovel  them 
out,  as  they  are  non-conductors  of  heat  which  effectually  prevent 
warming  up.  The  slag-spout  should  be  plugged  with  clay  so  as 
to  hold  the  slag  back  in  the  furnace.  When,  by  looking  in  at  the 
tuyeres,  we  see  the  slag  about  at  their  level,  tap  off  the  slag, 
which  should  quickly  fill  the  forehearth.  The  matte,  arising  from 
the  ore  charge,  now  begins  to  accumulate  in  the  forehearth,  and 
whenever  there  is  enough  for  a  tapping,  it  is  removed,  generally 
into  suitable  pots.  The  slag,  as  it  flows  from  the  forehearth,  may 
be  caught  in  slag-pots  or  it  may  be  granulated.  In  charging,  care 
should  be  taken  to  have  an  even  distribution  of  the  charge,  not 
coarse  ore  in  one  place  and  fine  in  another.  Unless  particular  care 
is  taken  in  this  respect,  we  may  have  irregular  running,  blast  and 
flame  coming  up  in  one  place,  and  the  furnace  apparently  dead  in 
another.  Two  methods  of  removing  the  slag  and  matte  from  the 
furnace  prevail,  intermittent  or  continuous  flow.  In  the  first 
method  the  slag  is  allowed  to  accumulate  on  the  furnace  hearth 
and,  before  it  rises  as  high  as  the  tuyeres,  is  tapped  off  by  pierc- 
ing the  clay  plug  with  which  the  tap-hole  has  been  closed.  This 
tap-hole,  on  account  of  the  corrosive  nature  of  the  matte,  is 


OF   THE    COMMON    METALS.  261 

water-cooled,  being  then  called  a  tap-jacket.  This  jacket  is  se- 
cured to  the  front  of  the  furnace,  and  is  about  10  in.  square.  By 
the  other  method  of  continuous  flow,  or  'open  breast/  the  slag 
and  matte  continuously  flow  from  the  furnace,  as  it  forms, 
through  a  trapped  spout,  as  shown  in  the  description  of  the  mat- 
ting furnace.  This  method  needs  less  attention  and  gives  less 
trouble  than  that  of  intermittent  tapping. 
Blast-funmce  copper-matte  smelting  of  roasted  ore. 

Roasted  ore,  together  with  oxidized  or  silicious  ores,  and  fuel, 
may  be  smelted  with  such  fluxes  as  may  be  necessary  to  produce 
a  suitable  slag.  Taking  the  case  of  a  furnace  in  full  operation, 
we  find  it  filled  with  charges  to  the  feed  door,  or  to  the  depth  of 
say,  7  ft.  As  the  molten  materials  are  drawn  off  below,  the  sur- 
face of  the  charge  imperceptibly  sinks,  making  room  for  more 
material.  The  coke  is  spread  out  in  a  layer  over  this  surface,  and 
upon  that  is  spread  the  calculated  and  weighed  charge,  the  fur- 
nace being  kept  full.  The  air  from  the  blowers  blown  into  the 
furnace  under  the  pressure  of  12  oz.  to  the  sq.  in.  burns  the 
coke  mostly  at  the  tuyeres.  The  gases  resulting,  together  with 
the  fumes  from  the  volatilization  of  a  portion  of  the  sulphur  pres- 
ent, come  up  through  the  charge  as  a  whitish  smoke  which  passes 
off  at  the  stack.  Most  of  the  sulphur  present,  uniting  itself  to  the 
copper  and  to  a  part  of  the  iron,  forms  a  matte  of  copper-iron 
sulphide.  This  matte,  when  forming,  will  also  take  up  the  con- 
tained gold  and  silver  existing  in  the  ore.  The  glowing  coke  of 
the  charge  reduces  the  iron  oxide  (unused  by  the  matte)  to  the 
ferrous  form,  and  this,  together  with  the  CaO,  A12O3  and  other 
bases,  fuses  with  the  silica  to  form  a  slag,  fluid  at  the  high  tem- 
perature at  the  tuyeres.  This  molten  material,  both  slag  and 
matte,  flows  from  the  furnace  into  a  forehearth,  a  box  or  tank 
in  which,  owing  to  the  difference  in  specific  gravity,  the  matte 
settles  out  to  the  bottom.  The  supernatant  slag  thus  freed  from 
the  matte,  escapes  by  an  overflow  spout  on  the  opposite  or  front- 
side  of  the  forehearth.  Here  it  may  be  received  into  slag  pots  to 
be  drawn  away  to  the  dump,  or  the  falling  stream  of  slag  may  be 
granulated  by  a  jet  of  water  which  sweeps  it  away,  carrying  it 
by  an  iron-lined  launder  to  the  dump.  The  matte  is  tapped  off 
from  the  forehearth  from  time  to  time,  as  it  accumulates,  into 


262  THE    METALLURGY 

pots.  Otherwise  it  may  be  tapped  into  a  sand-bed  in  which  are 
molded  the  depressions  for  the  pigs,  these  depressions  being 
joined  by  a  notch  from  one  to  another,  so  that  the  overflow  of  the 
first  molds  goes  into  the  succeeding  ones. 


72.     PYRITIC  MATTE  SMELTING. 

This  operation  is  performed  upon  iron-sulphide  ores  more  or 
less  silicious,  containing  some  copper  and  also  values  in  gold  and 
silver,  which  are  to  be  recovered.  It  consists  of  smelting  these 
ores  in  a  blast-furnace  with  a  small  amount  of  coke, 
burning  off,  say,  70  to  80%  of  the  sulphur  by  means  of 
the  air-blast,  while  the  remainder,  uniting  itself  to  the 
copper  and  iron  present,  forms  a  matte  which  takes  up 
the  contained  precious  metals  of  the  ores.  A  slag  is  formed 
from  the  silica  of  the  ore  gangue  and  unites  itself  to  the  bases 
present,  including  those  which  have  been  added  as  flux.  The 
matte  and  the  slag,  as  they  flow  from  the  furnace,  separate  in  the 
forehearth  according  to  specific  gravity,  the  slag  then  flowing 
away,  while  the  separated  matte  is  retained. 

Much  of  the  ore  contains  pyrite  (FeS2),  and  its  first  equivalent 
of  sulphur  is  easily  driven  off  by  the  heat  of  the  upper  part  of  the 
charge,  leaving  FeS.  This  is  acted  on  by  the  rising  air  blast  with 
the  development  of  heat  thus 

FeS  +  3  O  =  FeO  +  SO2 
22800  66400  +  71000=  113600 

We  have  heat  developed  by  the  burning  of  the  iron  and  sulphur 
equal  to  137,000  calories,  while  that  used  up  in  the  decomposition 
of  the  FeS  equals  22,800.  The  difference,  113,000  calories,  is  the 
resultant  heat  of  the  reaction.  Dividing  113,000  cal.  by  88,  the 
equivalent  of  FeS,  we  obtain  1,300  Ib.  calories  as  the  heat  devel- 
oped by  the  burning  of  one  pound  of  FeS. 

It  will  be  noticed  that  the  slow  and  expensive  preliminary  roast- 
ing of  the  sulphide  ores  is  dispensed  with  and  that  the  amount  of 
fuel  is  small  (2  to  6%),  because  of  the  heat  developed  by  the 
burning  of  the  pyrite  of  the  charge.  On  the  other  hand,  it  has 
been  found  to  be  an  advantage  to  use  hot  blast,  which  increases 


OF    THE    COMMON    METALS.  263 

the  intensity  of  combustion  at  the  tuyere-zone,  where  it  is  needed 
for  the  complete  and  thorough  fusion  of  the  slag,  and  for  vig- 
orous action  on  the  sulphides  of  the  charge.  The  expense  for 
fuel,  to  be  used  in  a  hot-blast  stove,  is,  however,  small  compared 
with  the  saving  of  coke  made  in  pyritic  smelting  over  that  in 
regular  matte-smelting.  Pyrite  ore  furnishes  a  considerable 
amount  of  iron,  which  is  available  for  fluxing  the  silica  in  iron- 
free  ores  that  .have  been  added  to  the  charge  for  the 
sake  of  recovering  their  contained  gold  and  silver  values. 
Where  necessary  limestone  and  iron  ore,  generally  the 
former,  are  added  as  flux.  An  iron  matte  alone  will 
not  remove  all  the  metals  from  the  ore,  and  it  has  been 
found  that  copper  to  the  extent  of  0.5%,  and  preferably  more, 
should  be  present  in  the  charge  to  ensure  the  collection  of  the 
values  in  the  matte.  The  slag  from  such  an  operation  will  then  be 
nearly  free  from  silver  and  gold.  Copper  therefore  acts  inci- 
dentally as  a  collector  of  values  of  gold  and  silver  in  the  charge. 
The  result  of  burning  off  so  much  sulphur  from  the  ore  to  the 
extent  of  70  to  80%  of  its  total  weight  means  that  only  the  30  to 
20%  left  is  available  to  form  matte,  first  with  the  copper  and  then 
the  iron,  in  accordance  with  its  needs.  It  is  this  removal,  in  the 
furnace,  of  such  large  amounts  of  sulphur  which  enables  one  to 
dispense  with  roasting  and  to  diminish  the  quantity  of  matte 
(matte  fall),  so  that  the  concentration,  or  ratio  of  charge  to 
matte,  is  large.  Thus,  suppose  a  pyrite  charge  contains  30% 
S,  and  that  its  volatilization  is  So% .  The  remaining  6%  of  the 
original  sulphur  would  make  24%  of  matte,  the  matte  containing 
25%  of  sulphur.  This  would  mean  a  concentration  of  100%  ore 
into  24%  matte,  or  of  about  4  into  I.  Again  suppose  we  have  a 
roasted  charge,  containing  10%  sulphur  only,  but  in  which,  as  is 
the  case  in  a  well  roasted  charge,  the  sulphur  loss  is  only  20%. 
Then  there  will  be  again  8%  of  sulphur  left  for  the  formation  of 
matte,  or  as  before,  24^  on  the  charge,  which  is  the  same  degree 
of  concentration  as  before.  So  it  may  be  seen  that  the  percentage 
of  volatilization  of  the  sulphur  varies  with  the  nature  of  the 
charge,  being  low  when  using  only  roasted  and  oxidized  ores, 
and  high  where  raw  ores  are  used,  especially  pyrite,  where  the 
first  equivalent  of  sulphur  is  loosely  held. 


264  THE  METALLURGY 

73.     PYRITE  SMELTING  AT  LEADVILLE,  COLORADO. 

The  practice  at  the  Bimetallic  smelting  works  was  as  follows: 
The  object  was  to  collect  values  of  gold  and  silver  in  a  matte, 
low  grade  because  of  the  small  amount  of  copper  present  in  the 
charge.  This  matte  was  again  put  through  another  furnace  in 
order  to  concentrate  it  and  to  raise  its  grade. 

At  the  first  smelting  there  was  a  concentration  of  4  into  I.  The 
furnace  produced  slag  which  was  thrown  away,  and  matte  which 
was  sent  to  the  concentrating  furnace.  Here  the  matte,  together 
with  silicious  ore,  was  again  concentrated,  so  that  the  proportion 
of  charge  to  the  matte  produced  was  as  4  to  I,  or  of  low-grade 
matte  to  concentrated  matte,  as  3  to  I.  The  slag  from  this  furnace 
was  returned  to  the  ore  furnace  for  re-smelting,  since  it  contained 
enough  values  to  warrant  doing  so. 

74.     MATTE  CONCENTRATION. 

The  matte  from  the  blast-furnace  smelting  of  pyrite  ore  may 
be  as  low  as  10%  in  copper,  or  even  lower.  In  this  lies  one  of  the 
unconsidered  drawbacks  to  pyrite  smelting.  The  matte  is  too  low 
grade  in  copper  to  ship  away  at  once,  or  to  treat  for  copper;  it 
must  be  concentrated  to  a  higher  grade. 

Now  if  we  were  to  heap-roast  this  matte,  and  then  run  it  in  a 
blast-furnace  with  sulphur-free  silicious  ores,  we  would  certainly 
get  the  copper  in  a  much  lessened  quantity  of  matte.  There  is, 
however,  the  expense  and  delay  of  this  roasting  to  consider,  and  it 
has  been  sought  to  smelt  the  matte  raw,  together  with  silicious 
ore,  with  the  idea  of  burning  the  sulphur  from  the  matte  so  as 
to  produce  a  concentration.  In  regular  matte  smelting,  if  this 
were  attempted,  the  matte  would  run  through  undiminished  in 
quantity ;  but,  by  the  newer  method,  having  a  limited  fuel  and  an 
abundant  blast,  some  concentration  has  been  obtained,  as  men- 
tioned in  the  description  of  pyrite  smelting  as  practised  at  Lead- 
ville,  Colo.  The  grade  of  shipping  matte  should  be  at  least  50% 
in  copper,  when  it  may  be  at  once  converted  to  blister  copper  by 
bessemerizmg.  However,  where  this  is  difficult  to  attain,  lower 
grade  mattes  are  often  made  and  shipped. 


OF    THE    COMMON    METALS. 


Slag  disposal  in  copper  blast-furnace  smelting. — The  slag 
from  a  blast-furnace,  being  a  waste  material,  is  gotten  rid  of 
in  the  cheapest  manner  possible.  In  the  case  of  the  smaller 
furnaces,  as  it  flows  from  the  forehearth,  it  is  caught  in  wheeled 
slag-pots  (called  slag-carts).  These  are  taken  by  men  to  the 
edge  of  the  dump  and  there  poured  out.  As  the  dump  in- 
creases in  length,  however,  this  becomes  an  increasing  expense, 


FIG.  122.     BENNETT'S  SLAG-CASTING  MACHINE. 


and  then  large  slag  cars  are  used,  which  run  on  a  track  from  the 
furnace  to  the  edge  of  the  dump.  These  cars  are  operated,  either 
by  horsepower  or  by  an  industrial  electric  locomotive.  Another 
favorite  way  is  to  granulate  the  slag.  For  this  a  cast-iron 
launder  is  arranged  to  receive  it,  and  in  the  launder  is  a  good  flow 
of  water.  In  addition  a  horizontal  flattened  jet  of  water  is  ar- 
ranged to  strike  the  falling  slag,  cooling  and  breaking  it  up,  when 
it  is  swept  away  in  the  launder  to  the  dump.  The  launder  should 
have  a  grade  of  one  inch  to  the  foot. 


266  THE  METALLURGY 

75.     HEAT  EQUATIONS  OF  THE  COPPER-MATTING  FURNACE. 

In  smelting  raw  sulphide  ores  containing  iron  pyrite  (FeS2) 
or  pyrrhotite  (FeS),  together  with  copper  sulphides  (CuS), 
they  are  acted  on  by  the  heat  to  drive  off  the  first  equivalent  of 
sulphur  from  the  FeS2,  leaving  FeS.  The  air  of  the  blast  acting 
on  this  then  oxidizes  it  according  to  the  equation 

FeS  +  30  =   FeO  +  SO2 
23800       .          66400         71000=113600 

Now  it  is  to  be  noticed  that  where  copper  is  present  in  the 
charge  it  retains  sulphur  which  also  takes  up  iron  as  FeS  to  form 
a  mixed  sulphide  or  matte  CuFeS.  Analyses  of  matte  show  that 
the  sulphur  varies  from  20  to  26%  (average  23%),  and  that  the 
iron  and  copper  together  varies  from  65  to  72%  (average  68%). 
These  approximate  figures  are  useful  in  calculating  a  charge. 

76.     THE  REVERBERATORY  MATTING  FURNACE. 

Fig.  115  represents  in  elevation  and  Fig.  116  in  sectional  plan, 
a  large  reverberatory  matting-furnace,  having  a  hearth  37.5  ft. 
long  by  15  ft.  wide,  with  a  fire-box  7  by  8  ft.,  or  56  sq.  ft.  area. 
There  are  six  hoppers  for  charging  ore  and  one  double  hopper 
by  which  coal  can  be  charged,  as  needed,  to  the  fire-box.  The 
fire  is  'grated'  from  time  to  time,  when  the  doors  of  the  closed 
ash-pit  are  opened,  and  men  enter  to  remove  the  ashes  and  clink- 
ers from  between  the  grate  bars  by  means  of  a  long-handled 
pointed  bar.  This  material  falls  into  cars  set  on  a  track  shown 
as  entering  the  ash-pit  (see  Fig.  116).  The  fire  having  been 
cleaned,  the  ash-pit  doors  are  closed  and  air  is  forced  into  it 
under  low  pressure  by  a  pipe  (shown  on  the  plan)  next  to  the 
fire-box.  There  are  four  side-doors,  two  for  skimming  at  the 
near  side,  and  one  front  door  where  the  slag  is  also  skimmed. 
The  stack  is  separate  from  the  furnace,  the  smoke  escaping  to  it 
by  means  of  a  sloping  flue  from  the  outlet-port  in  the  roof  of  the 
furnace.  On  the  far  side,  between  the  two  side  doors,  may  be 
seen  the  matte  tap-hole  set  at  the  level  of  the  bottom  of  the  hearth, 
from  which  the  matte  is  withdrawn  as  it  accumulates. 


OF    THE    COMMON    METALS. 


267 


268  THE  METALLURGY 

77.     REVERBERATORY  MATTE  SMELTING. 

While  the  blast-furnace  in  general  affords  the  cheapest  means 
of  smelting  copper-bearing  ores  in  lump  or  coarse  form,  one  ob- 
jection to  it  is  that  the  strong  blast  which  must  be  used  sweeps 
away  a  good  deal  of  the  fine  and  dusty  ore,  forming  flue-dust.  It 
is  true  that  such  flue-dust  may  be  in  large  part  caught  in  flues,  and 
made  into  briquettes  and  re-smelted,  but  it  is  always  an  addi- 
tional expense  which  is  to  be  avoided.  In  fact,  fine  ore  and  concen- 
trate are  better  treated  in  reverberatory  furnaces.  At  the  time  of 
dropping  such  a  charge,  the  stack  damper  may  be  closed,  and  when 
the  dust  has  settled  again  opened,  so  that  the  ore  is  not  dis- 
turbed until  it  is  melted.  Less  sulphur  is  volatilized  than  in  the 
blast-furnace,  and  the  ore  must,  therefore,  first  be  roasted.  The 
output  of  a  reverberatory  is  much  less  than  that  of  a  blast-furnace, 
and  there  i^  much  more  fuel  used,  though  of  a  cheaper  kind.  Thus 
one  would  use  in  reverberatory  smelting  21  to  33%  of  coal  as 
against  10%  coke  in  regular  blast-furnace  smelting,  or  as  low  as 
2  to  6%  in  pyrite  smelting.  We  might  say  that  the  25%  of  coal 
at  $3  per  ton  would  just  balance  the  10%  coke  at  $7.50  per  ton  so 
far  as  the  item  oi.fuel  is  concerned. 

Operation  of  the  furnace. — While  in  blast-furnace  work  the 
charges  are  put  into  the  furnace  continuously,  in  the  reverbera- 
tory, a  charge  of  many  tons  is  put  into  the  furnace  at  once.  The 
ore  is  stored  in  hoppers  situated  above  the  furnace,  into  which  it 
falls  when  the  hopper  slide  is  withdrawn.  Were  the  ore  to  drop 
directly  upon  the  floor  of  the  furnace,  the  cold  charge  would  tend 
to  stick  there  and  be  heated  with  difficulty.  It  is  customary, 
therefore,  to  retain  a  pool  or  layer  of  matte  on  the  hearth  upon 
which  the  ore  drops  and  spreads  out.  At  the  same  time  the  fire 
has  been  cleaned  or  'grated'  by  removing  clinkers  and  ashes.  The 
firing  now  proceeds  vigorously  until  the  charge  has  been  melted 
down.  The  sulphur  of  the  charge  unites  itself  to  the  copper  and 
to  the  iron  to  satisfy  its  needs,  and  the  matte  thus  formed,  sepa- 
rating out  in  drops,  takes  up  any  precious  metals  that  exist.  Then 
by  its  greater  specific  gravity  as  compared  with  the  slag,  it  finds 
its  way  to  the  hearth.  The  gangue  of  the  ore  yields  SiO2  to  unite 
with  FeO,  CaO,  A12O3,  and  with  minor  bases,  to  form  a  fusible 


OF    THE    COMMON    METALS. 
"I 


269 


FIG.  116.     REVERBERATORY  SMELTING  FURNACE  (PLAN). 


2/0  THE    METALLURGY 

slag,  which  floats  upon  the  matte  as  a  separate  layer.  The  side 
doors  of  the  furnaces  are  then  opened,  and  the  slag  skimmed  by 
means  of  long-handled  rabbles.  As  the  slag  is  removed  it  falls 
either  into  wheeled  slag-carts  or  into  pots  which  are  wheeled  away 
to  the  dump.  The  sides  of  the  interior  of  the  furnace,  where  they 
appear  to  have  been  eaten  out  or  eroded  by  the  action  of  the 
molten  charge  upon  them,  are  now  repaired  or  fettled  by  sand, 
placed  at  the  eroded  surface  with  a  long-handled  paddle,  or  by 
balls  of  ganister,  pressed  against  the  side  with  the  same  tool.  The 
side  doors  are  now  closed  and  luted  tight  with  a  stiff  clay  mortar, 
and  the  charge,  already  in  the  hopper,  is  dropped  in.  From  time 
to  time,  as  it  is  needed  or  as  it  accumulates,  the  matte  is  tapped 
off  at  a  tap-hole  situated  at  the  lowest  point  of  the  hearth.  The 
matte  runs  either  into  sand  molds  (depressions  made  by  means  of 
a  shovel  in  the  sand  floor  of  the  furnace-house)  or  into  ladles, 
operated  by  an  overhead  traveling  crane,  \vhere  it  is  to  be  trans- 
ferred to  a  converter. 

78.     REVERBERATOR v   SMELTING   OF   COPPER.      (WELSH    COPPER 

PROCESS.) 

The  process  consists  in  treating  copper  ores  by  a  series  of  roast- 
ings  and  fusions,  which  bring  up  the  grade  of  copper  contained 
in  the  charge  to  blister  copper,  which  is  subsequently  refined.  We 
may  divide  the  process  into  six  operations : 

1.  Roasting. — Sulphide  ores,  containing  copper  of  from  5  to 
15%,  are  roasted  in  a  hand  roaster,  until  only  5%  of  sulphur  is 
left. 

2.  Fusion. — These  roasted  ores  are  charged  into  a  reverber- 
atory  furnace  together  with  such  oxidized  ores  as  are  available, 
and  are  melted   down.     The   sulphur   contained   in  the   roasted 
ore  unites  itself  to  whatever  copper  is  present   (taking  up  also 
some  iron)    and  forms  a  matte  of  perhaps  35%   copper  called 
'coarse  metal.'    The  silica,  uniting  itself  to  the  remainder  of  the 
ferrous  or  ferric  oxide  and  to  alumina  and  other  bases,  forms  a 
slag,  fusible  at  the  high  heat  of  the  furnace.     The  slag  is  then 
skimmed  off  into  slag-pots  with  a  rabble,  which  is  inserted  at  the 
front-end  door.    The  remaining  matte  is  tapped  off  at  the  front- 


OF   THE    COMMON    METALS.  2/1 

side  of  the  furnace  into  a  channel  or  runner  arranged  in  the  sand 
floor,  and  from  this  into  cavities  in  the  sand  to  form  pigs.  This 
operation  resembles  that  performed  at  an  iron  blast-furnace  in 
casting  pig  iron.  The  flow  of  matte  is  stopped  before  the  furnace 
is  quite  empty,  some  slag  and  matte  still  remaining,  and  another 
charge  is  dropped  in  from  the  hoppers.  Such  a  charge  takes  six 
hours  to  work  through. 

3.  The  coarse  metal   (matte)   from  the  sand  beds  is  crushed 
to  5-mesh  and  passed  on  to  another  roasting  furnace.     Whatever 
rich  sulphides  of  from  20  to  70%  Cu  are  on  hand  are  also  charged 
in,  and  with  the  matte  roasted  to  5%  sulphur. 

4.  This  roasted  material  is  charged  to  a  reveberatory  fusion 
furnace,  together  with  oxidized  ores  containing  from  20  to  70% 
copper.     When  the  charge  has  been  melted  there  results  a  matte 
called  'white  metal,'  containing  75%  copper  and  little  else  than 
copper  and  sulphur.     As  before,  the  silica  from  the  oxidized  ore 
unites  with  the  ferrous  iron  and  other  bases  to  form  a  slag.    This 
slag,  however,  having  been  made  from  such  rich  materials,  con- 
tains too  much  copper  to  be  rejected,  and  is  turned  back  to  the 
fusion  furnace  described  in  operation  2. 

5.  The  white  metal  in  pigs  is  charged  into  a  reverberatory  fu- 
sion furnace  where  the  pigs  are  piled  up  in  open  order.    The  fire 
is  gradually  increased  so  that  it  takes  several  hours  to  melt  them 
down.    This  operation  is  called  'roasting.'    The  highly  heated  air 
admitted  at  the  fire-box  acts  on  the  surfaces  of  the  pigs  and  upon 
the  drops  of  matte  trickling  down  from  them,  converting  a  portion 
into  cuprous  oxide.     We  thus  get  copper  present  both  as  oxide 
and  as  sulphide.     Finally,  the  heat  is  increased  until  the  whole  is 
melted  down.    Now  occurs  the  endothermic  reaction, 

2Cu2O  +  CuS  =  3Cu  +  SO, 

2X420OO   202OO  7IOOO  =  3320O 

in  which  copious  fumes,  boiling  up  from  the  surface  of  the  molten 
charge,  are  given  off.  Some  slag,  rich  in  copper,  is  also  produced, 
getting  its  silica  partly  from  the  borders  of  the  furnace  and  partly 
from  some  silicious  but  perfectly  oxidized  ore  which  has  been 
added  to  the  charge.  This  slag  is  returned  to  operation  4.  Fi- 
nally, there  is  tapped  off  blister  copper,  so  called  because  when 


272  THE    METALLURGY 

cooling,  occluded  gas  seeking  to  escape  forms  blisters  on  the  sur- 
face of  the  pigs  of  metal. 

6.  The  blister  copper  now  contains  98%  copper,  but  also  im- 
purities, which  must  be  removed  before  it  is  fit  for  market.  This 
refining  process  is  elsewhere  described.  In  case  the  copper  con- 
tains gold  and  silver,  taken  up  from  the  ores  from  which  it  has 
been  made,  it  is  customary  to  re-melt  it,  to  pole  it  to  remove  cop- 
per oxide,  and  to  cast  it  into  anodes  for  electrolytic  refining. 

\Yhere  copper  ores  are  impure,  or  where  it  is  desired  to  obtain 
a  superior  grade  of  copper  (called  'best  selected'),  the  process 
is  varied  as  follows:  A  portion  of  the  75%  white  metal  pro- 
duced in  operation  4  is  ground  up  and  roasted,  thus  producing 
Cu2O,  together  with  some  Cu2S,  due  to  sulphur  still  unremoved. 
This  roasted  matte  is  put  in  a  fusion  furnace  and  promptly  melted 
and  stirred  in,  together  with  lump  white  metal.  A  reaction  sets 
in,  such  as  was  described  in  operation  5,  a  small  portion  of  the 
copper  reducing  out  from  the  Cu2O,  and  in  so  doing  picking  up 
the  impurities  present,  such  as  As,  Sb  Te,  and  also  contained 
gold.  On  the  other  hand  the  white  metal  which  remains  is 
quite  pure  and  is  capable  of  giving  a  pure  grade  of  copper  when 
passed  on  to  operation  5.  After  skimming  the  charge,  the  matte 
is  tapped  out  as  usual  into  sand  molds.  In  the  first  few  molds  will 
be  found  the  copper  which  has  been  reduced,  lying  at  the  bottom 
beneath  the  matte  which  forms  the  body  of  the  pig — whence  the 
name,  'bottoms.'  These  bottoms  can  be  re-melted  and  cast  into 
anodes  for  the  electrolytic  refinery,  where  the  precious  metals 
can  be  conveniently  extracted. 

79.     DIRECT  TREATMENT  OF  COPPER  MATTE  FOR  BLISTER  COPPER. 

In  this  modification  of  the  Welsh  method  of  producing  blister 
copper,  instead  of  'roasting'  the  matte  in  lump  form  in  the  blister 
furnace,  that  operation  is  performed  in  a  separate  roasting  fur- 
nace either  of  the  automatic  or  of  the  reverberatory  type.  The 
charge  for  the  blister  furnace  will  then  consist  of  14,000  Ib.  of 
75%  roasted  'white  metal'  or  matte,  still  retaining  4  to  6%  of 
sulphur,  3,500  Ib.  raw  white  metal  or  matte,  4,000  to  8,000  Ib.  of 
foul  slag  from  a  former  charge  and  600  Ib.  of  silicious  ore  to  take 


OF    THE    COMMON    METALS.  273 

up  the  FeO  and  other  bases  present.  When  this  has  been  melted 
down,  6,000  Ib.  more  of  roasted  matte  is  added,  making  the  total 
of  matte  charged  23,500  Ib.  and,  when  fused,  the  reaction  begins. 
The  surface  of  the  charge  is  seen  to  be  seething  with  escaping 
bubbles  of  gas  which  are  set  free  according  to  the  reaction : 

Cu2S  +  2  Cu2O  =  6  Cu  +  SO2 
20200     2X42000  71000  =  — 33200  pound  calories. 

The  reaction  is  endothermic  requiring  87  calories  per  pound  of 
copper  produced.  The  bath  is  then  skimmed  to  remove  the  slag, 
which  contains  considerable  copper  oxide,  and  the  blister  copper 
is  ready  to  be  ladled  into  pigs.  The  method  has  the  advantage 
that  the  output  is  greater  than  by  the  old  method,  and  the  quality 
of  the  copper  quite  as  good.  From  the  above  charge  there  is  pro- 
duced 75  pigs  of  230  Ib.  each,  or  17,250  Ib.  of  blister  copper,  and 
23  pots  slag  of  400  Ib.  or  9,200  Ib.  containing  12  to  15%  Cu. 
Where  'best  selected  copper'  is  made  there  is  left  behind  a  quan- 
tity of  bottoms  containing  impurities.  These  bottoms  may  be 
worked  up  as  follows :  A  charge  is  put  in  the  blister  furnace  con- 
sisting of  14,000  Ib.  of  roasted  75%  matte,  21,000  Ib.  of  raw  matte 
of  the  same  grade,  8,000  Ib.  of  bottoms  and  1,000  Ib.  of  silicious 
ore.  When  melted  this  is  followed  by  6,000  Ib.  of  75%  roasted 
matte,  the  whole  charge  being  50,000  Ib.  This  is  treated  precisely 
like  the  former  charge,  but  it  gives  a  much  higher  yield  of  copper. 

80.    REVERBERATORY  MATTE-SMELTING  AT  ANACONDA,  MONTANA. 

Fig.  117  and  118  represent  in  elevation  and  plan  the  No.  i  fur- 
nace at  the  Washoe  plant  of  the  Anaconda  Copper  Mining  Co., 
with  two  Sterling  boilers  heated  by  the  waste  gases  from  the  fur- 
nace and  developing  600  h.  p.  The  furnace  will  treat  on  an  average 
275  tons  in  24  hours,  producing  a  40%  matte  with  a  concentration 
of  4  into  i.  The  charge,  consisting  of  hot  calcines  from  the 
McDougal  roasters,  has  by  analysis  9%  Cu,  24.4%  FeO,  2.9% 
CaO,  8.1%  S,  and  26%  SiO2.  Every  80  minutes  a  charge  of  15 
tons  is  dropped  into  the  furnace  from  the  two  hoppers  nearest  the 
fire-bridge  (shown  in  full  lines  in  the  elevation).  This  falls  upon 
the  bath  of  molten  matte  and  slag.  Being  hot,  it  spreads  out  in  all 


2J4  THE    METALLURGY 

directions  and  floats  slowly  down  the  furnace  toward  the  front. 
It  is  readily  melted  by  contact  with  the  molten  slag  and  matte 
bath  below,  and  the  flame  above  it.  In  this  great  reservoir  of  heat 
there  is  but  little  variation  of  temperature,  and  the  flame  is  trans- 
parent. Because  of  the  width  of  the  furnace,  the  sides  are  less 
corroded  than  in  a  narrower  one.  The  fuel  is  soft  coal,  56  tons 
or  21%  of  the  charge  being  burned  daily  upon  110.7  scl-  ft-  or 
grate  area,  or  40  Ib.  per  sq.  ft.  hourly.  Every  four  hours  45  to  50 
tons  of  slag  is  in  15  minutes  removed  from  the  furnace,  it  being 
allowed  to  flow  off  by  the  front  door  in  a  thick  stream,  which,  as 
it  falls  toward  the  waste  launder,  is  granulated  by  a  strong  hori- 
zontal stream  of  water,  which  also  sweeps  it  away  to  the  dump, 
situated  at  some  distance  from  the  furnace.  The  matte  is 
kept  at  nearly  the  same  level,  some  10  tons  being  tapped  out 
at  a  time,  while  the  total  amount  of  it  in  the  furnace  may  be 
one  or  two  hundred  tons.  The  action  of  the  slag  upon  the  fur- 
nace is  to  corrode  or  scour  it,  especially  at  the  fire-bridge.  It 
must  therefore  be  repaired.  This  is  done  monthly  by  drawing  off 
the  slag  and  matte  completely  and  throwing  in  sand  or  silicious 
ore  against  the  corroded  sides  until  they  have  been  effectually 
banked  up.  One  of  these  furnaces  will  run  for  six  or  eight  months 
before  it  has  to  be  shut  down  for  renewal  of  the  corroded  parts 
and  repair  of  the  roof  and  sides.  The  exposed  inner  surfaces  of 
the  furnace  are  composed  of  silica  brick  which  is  practically  infu- 
sible but  which  expands  in  heating  so  that  an  allowance  has  to  be 
made  for  it  by  leaving  transverse  slits  in  the  roof.  These  close  up 
when  the  brickwork  is  at  its  full  temperature.  An  important  point 
in  effectual  working  is  the  outlet  flue  or  neck,  60  by  38  in.  in  di- 
mensions or  16 sq.ft. area,  which  must  be  large  enough  to  produce 
the  draft  and  yet  hold  back  the  flame  through  the  furnace.  At  the 
fire-box  the  ash  and  cinder,  as  they  fall,  drop  into  a  stream  of 
water  which  carries  them  away.  Hence  this  place  is  always  cool 
and  accessible,  and  the  fire  is  kept  constantly  clean  and  properly 
'grated.'  A  good  deal  of  coke  or  cinder  from  the  coal  falls  down, 
and  this  is  later  concentrated  to  remove  the  ashes  and  clinkers. 
This  residue,  still  having  a  good  deal  of  heat  value,  is  used  to  mix 
in  with  the  flue-dust  in  the  making  of  briquettes  for  the  blast- 
furnace. 


OF   THE    COMMON    METALS. 


275 


276  THE  METALLURGY 

81.     HENDERSON  PROCESS  FOR  EXTRACTION  OF  COPPER  FROM 
BURNT  PYRITE. 

The  residue  or  cinder  left  from  pyrite  used  in  sulphuric  acid 
making  may  contain  from  2  to  4^  Cu,  together  with  some  silver 
and  gold;  and  all  these  metals  can  be  extracted  from  it  by  a 
chloridizing  roast,  followed  by  a  leaching  with  water  and  a  pre- 
cipitation with  scrap-iron. 

Fig.  119  gives  a  plan  and  transverse  sectional  elevation  of  a 
200- ton  plant  of  the  Pennsylvania  Salt  Manufacturing  Co.,  Na- 
trona,  Pa.  It  is  operated  as  follows. 

The  cinders  (roasted  ore)  are  ground  dry  in  a  pan-mill,  simi- 
lar to  a  Chilean  mill,  to  2O-mesh,  and  mixed,  during  the  opera- 
tion, with  12%  of  their  weight  of  salt.  The  mixture  goes  by  belt- 
elevator  to  storage  bins  commanding  the  charge-floor,  k,  of  the 
roasting  furnaces,  b.  It  is  weighted  in  charges  of  5  tons  to  the 
charge-tubes  or  hoppers  of  the  roasters,  represented  in  the  longi- 
tudinal sectional  elevation  in  Fig.  120.  The  four  2O-in.  charge- 
tubes  are  clearly  shown. 

The  gases  from  the  fire  pass  along  the  14-in.  space  above  the 
main  furnace  or  muffle,  and  going  down  the  flue  at  the  end,  return 
under  the  8  by  35  ft.  hearth  of  the  muffle.  Finally,  they  pass  by  a 
main  underground  flue,  2  ft.  by  2  ft.  in  size,  to  a  common  stack. 
The  gases  from  the  roasting  ore  pass  by  an  i8-in.  pipe  to  the  con- 
densing towers,  a,  where  they  pass  upward  through  coke  in  lumps. 
Water  trickles  down  through  this  coke  from  above,  absorbing  the 
chlorine  and  the  hydrochloric  acid  developed  from  the  gases.  For 
such  a  chloridizing  roast,  muffle  furnaces  have  superseded  the 
reverberatory,  since  the  former  kind  has  the  advantage  of  need- 
ing but  one-half  of  the  condensing  capacity,  and  better  results  are 
obtained,  due  to  the  separate  control  of  the  fire  and  of  the  gases 
of  the  reaction. 

Enough  raw  pyrite  is  charged  with  the  ground  cinder  to  bring 
the  proportion  of  sulphur  to  1.5  times  the  contained  copper.  The 
whole  is  brought  to  a  just-visible  red  heat  (525°  C)  and  kept  well 
stirred.  The  total  time  for  a  charge  is  8  hours.  The  finished 
charge  is  drawn  upon  the  floor,  allowed  to  cool,  shoveled  into 
charge-cars,  elevated  to  the  charge-floor  level,  and  delivered  to 


OF    THE    COMMON    METALS. 


277 


2/8  THE    METALLURGY 

the  leaching  tanks,  d,  12  by  12  by  4  ft.  in  size.  These  tanks  are 
made  of  three-inch  plank  filled  in  the  3-in.  space  between  the  inner 
lining  and  the  outer  shell  with  a  mixture  of  i  to  i  sand  and  tar. 
The  false  bottom  of  the  leaching  tank  is  made  of  a  paving  of 
brick  covered  with  small  pieces  of  refuse  coke.  The  calcined  ore 
is  first  lixiviated  by  weak  liquor  from  a  previous  operation,  then 
by  water,  and  finally  by  the  weak  hydrochloric  acid  from  the  con- 
densing towers,  all  being  run  to  the  settling  tanks,  c,  12  by  12  by  6 
ft.  in  size.  The  first  removes  the  bulk  of  the  copper,  becoming  a 
strong  solution,  the  second  becomes  the  weak  solution  for  the 
succeeding  operation,  and  the  third  dissolves  cupric  oxide  and 
cuprous  chloride  otherwise  insoluble.  The  residue  (purple  ore) 
from  the  leaching  vats  is  shoveled  out  upon  the  floor  c,  thence  to 
be  loaded  upon  railroad  cars  set  below.  The  weak  liquors  are 
pumped  back  to  the  lixiviation  tanks,  and  the  strong,  when  they 
have  reached  18°  Beaume,  are  drawn  off  to  the  tanks,  /  (12  by  12 
bv  6  ft.),  where  the  contained  copper  is  precipitated  by  means  of 
thin  scrap-iron.  The  tanks  have  open-slat  false  bottoms  about  2 
ft.  above  the  real  bottom.  Live  steam,  blown  into  the  solution, 
agitates  it,  and  the  copper  precipitate  works  down  between  the 
slats  upon  the  real  bottom,  and  is  removed  to  the  tanks,  g,  10  by 
10  by  5  ft.  The  chloride  solutions  from  this  tank  flow  over  scrap- 
iron  placed  in  the  small  vats,  h,  containing  scrap-iron,  which 
catches  any  escaping  particles  of  precipitate.  The  latter  contains 
90%  Cu,  35  oz.  Ag  and  0.15  oz.  Au  per  ton. 

Before  precipitating  the  copper  it  would  be  possible  to  remove 
the  silver  by  the  Claudet  process,  precipitating  it  by  means  of  zinc 
iodide.  It  has  been  found,  however,  more  profitable  to  sell  it  di- 
rect to  the  blue  vitriol  makers,  who  pay  for  95%  of  the  silver  and 
for  the  full  value  of  the  contained  copper  and  gold. 

The  cost  of  treatment  by  the  Henderson  process,  with  common 
labor  at  $1.50  per  day,  is  $1.87  per  ton. 

82.     THE  HYDROMETALLURGY  OF  COPPER. 

Extraction  of  copper  b\  wet  methods. — This  consists  in  getting 
the  copper  into  aqueous  solution,  with  or  without  the  aid  of  sol- 
vents, the  copper  being  in  combinations  suitable  for  solution.  It 


OF    THE    COMMON    METALS. 


279 


28O  THE    METALLURGY 

is  then  precipitated  by  metallic  iron  (iron  scrap)  or  other  precipi- 
tant, the  precipitate,  contaminated  by  small  pieces  of  scrap-iron 
and  iron  rust  (iron  salts)  being  melted  down  and  refined  in  a 
reverberatory  furnace.  In  order  to  extract  copper  from  the  ore, 
the  latter  must  be  put  in  form  to  be  acted  upon,  either  by  crush- 
ing or  by  roasting. 

Copper  can  be  profitably  extracted  from  suitable  low-grade 
copper-bearing  ores  of  as  low  as  0.5  to  i%  Cu,  especially  where 
labor  is  cheap. 

Ores  containing  quantities  of  lime,  magnesia,  ferrous  oxide  or 
manganese  are,  however,  not  suited  for  wet  processes.  Ores  con- 
taining poor  oxides,  carbonates,  or  sulphates,  and  having  a  quartz 
gangue,  are  the  best.  Copper  sulphides  should,  however,  be  added 
to  this  list,  since  they  may  be  converted  into  sulphates,  either  (i) 
by  natural  decomposition,  or  (2)  by  a  slow  roasting  of  the  ore. 
Copper  sulphides  can  also  be  converted  into  cuprous  or  cupric 
chlorides  which,  again,  are  soluble  in  common  salt.  If  the  ore 
contains  limestone,  then  by  roasting  in  a  kiln,  caustic  lime  can  be 
produced,  capable  of  being  leached  out  by  water.  Afterward  the 
ore  can  be  treated  for  the  extraction  of  the  copper. 

Extraction  of  copper,  rendered  soluble  by  natural  decomposi- 
tion.— In  mines,  a  gradual  oxidation,  due  to  the  action  of  air  and 
moisture  goes  on,  so  that  mine  water  becomes  impregnated  with 
copper  sulphate  which  gives  it  a  blue  tint.  Such  water,  run 
through  boxes  or  launders  containing  scrap-iron,  precipitates  out 
its  contained  copper,  which  replaces  the  iron  of  the  scrap.  Thus 
we  have  a  source  of  revenue.  At  Jerome,  Arizona,  such  waters 
are  run  through  extended  troughs  or  launders  containing  scrap- 
iron. 

Pyrite  leaching  at  Rio  Tinto,  Spain. 

This  well-known  method,  adopted  for  the  extraction  of  copper, 
consists  in  allowing  huge  heaps  of  the  mineral  to  oxidize  under 
the  influence  of  air  and  moisture,  and,  subsequently  washing  out 
the  copper  sulphate  as  it  is  formed  by  running  water 
through  the  heap.  The  copper  is  precipitated  from  this  solution 
by  pig  iron.  When  the  copper  occurs  as  chalcopyrite,  CuFeS2, 
or  as  covellite,  CuS,  oxidation  proceeds  slowly.  The  best  form 


OF    THE    COMMON    METALS.  28l 

is  chalcocite  or  copper  glance,  Cu2S,  which,  with  pyrite,  FeS2, 
constitues  the  bulk  of  Rio  Tinto  sulphide.  When  the  mineral 
is  subjected  to  the  combined  influence  of  air,  heat,  and  moisture 
in  the  heaps,  the  following  reactions  take  place : 

!.  FeS2+7O+H2O=FeSO4+H2SO4,  and  this  ferrous  sul- 
phate easily  oxidizes  to  ferric  sulphate. 

2.  2  FeSO4+H2SO4+O=Fe2(SO4)8+H2O. 

This  ferric  sulphate  now  acts  on  the  copper  glance,  and  the 
contained  copper  is  rendered  soluble  thus : 

3.  Fe2  (SO4)3  +  Cu2S  =  CuSO4  +  2  FeSO4  +  CuS 
and  the  cupric  sulphide  is  changed  as  follows : 

4.  Fe2  (SO4)3  +  CuS  +3O  +  H,O  =  CuSO4  +  2  FeSO4  + 
H2S04. 

Reaction  3  is  fairly  rapid,  causing  about  half  the  copper  to  go 
into  solution  in  a  few  months,  reaction  4  is  much  slower,  and 
yields,  under  favorable  conditions,  80%  of  the  remaining  half  of 
the  copper. 

The  method  of  working  is  as  follows :  A  site  is  chosen  where 
the  ground  is  sufficiently  sloping  and  concave  to  enable  the  cop- 
per solution  to  collect  and  to  run  out  at  one  side  of  the  heap, 
which  may  contain  100,000  tons.  On  the  ground  is  first  arranged 
a  network  of  air  flues  12  in.  square,  made  of  rough  stones.  At 
,50  ft.  distant  from  one  another  vertical  chimneys,  also  of  rough 
stones,  connecting  with  the  flues,  are  built  up  as  the  heap  rises. 
The  ore  is  crushed  to  not  more  than  3  or  4  in.,  and  the  lump  and 
fine  are  separated.  Lump  ore  and  fine  are  alternately  dumped 
until  the  height  of  the  level-topped  pile  is  30  ft.  The  surface  of 
the  pile  is  formed  into  squares  by  "idges  of  fines.  The  squares 
are  to  ensure  more  even  distribution  of  the  water  to  the  heaps. 
Launders  are  also  provided  to  conduct  water  to  all  parts  of  the 
top  surface  of  the  heap.  As  the  heap  is  being  formed  water  is  run 
on  to  extract  any  already  existing  copper  sulphate.  Oxidation 
also  starts  as  the  result  of  the  wetting.  The  completed  and  wetted 
heap  begins  to  oxidize  pretty  rapidly,  as  shown  by  the  heat  pro- 
duced, the  temperature  in  the  chimneys  rising  to  80°  C.  As  the 
temperature  rises,  the  lower  surface  openings  should  be  closed  to 
control  oxidation  and  spread  it  through  the  heap.  The  surface  as- 
sumes a  brown  color  due  to  dehydration  of  the  basic  ferric  salt 


282  THE    METALLURGY 

which  forms,  and  the  gradual  heating  up  may  be  noted  by  this 
drying  action.  The  greatest  care  must  be  taken  to  prevent  the 
heap  from  firing.  When  oxidation  has  proceeded  as  far  as  safe, 
water  is  run  on  at  the  rate  of  220  gal.  per  minute  until  soluble 
copper  salts  are  leached  out.  The  heap  is  again  allowed  to  oxidize 
and  the  washing  out  repeated.  After  a  year  the  top  surface  needs 
're-tilling,'  the  ridges  are  arranged  where  the  squares  formerly 
were  and  the  launders  are  also  shifted.  At  the  edge  of  the  heap, 
for  a  distance  of  some  yards,  the  ore,  which  has  become  cemented, 
hold  much  copper  salts,  and  is  here  dug  down  into  terraces  in 
order  that  this  copper  may  also  be  extracted  by  washing.  When 
there  remains  but  0.3%  copper,  the  ore  may  be  called  leached. 

The  copper  liquor  which  runs  from  the  heap  contains  ferric 
oxide,  and  this  must  be  reduced  by  running  it  over  a  filter-bed  of 
fresh  mineral.  This  bed  is  placed  within  a  reservoir  formed  by 
a  masonry  dam  across  a  small  ravine.  The  liquor,  after  perco- 
lating the  bed,  remains  in  contact  with  it  until  drawn  off  to  the 
precipitating  tanks.  The  solution  contains  0.4%  Cu,  0.1%  Fe2O3, 
2%  FeO,  i%  H2SO4  and  0.03%  As.  The  large  quantity  of  FeO 
and  of  H2SO4  is  due  to  the  fact  that  a  part  of  the  waste  liquor 
from  the  precipitation  tanks  is  pumped  back  and  used  for  water- 
ing the  heaps,  so  that  the  solutions  tend  to  concentrate.  The 
liquor  at  the  reservoirs  is  run  through  precipitation  tanks  over 
pig  iron  piled  up  in  open  order  therein,  to  precipitate  the  copper 
in  the  form  of  'cement'  copper  or  'copper  precipitate.'  These  tanks 
are  arranged  on  the  slope  of  a  hill,  the  liquor  passing  back  and 
forth  through  the  tanks  until  discharged  from  the  lowest  tanks 
free  from  copper.  Some  of  the  tanks  are  cut  out  of  the  system, 
or  by-passed  daily,  the  liquor  meantime  going  through  the  re- 
maining tanks.  All  iron  is  removed  from  these  cut-out  tanks  and 
piled  up,  the  copper  on  the  iron  being  meanwhile  knocked  off  and 
thrown  back  in  the  tank.  The  dirty-looking  precipitate  is  now 
removed  to  the  cleaning  and  concentrating  plant,  while  the  iron 
is  piled  back  in  the  tank  and  the  liquor  again  turned  through  it. 

The  crude  precipitate,  containing  70%  copper,  is  by  means  of  a 
strong  jet  of  water  gradually  worked  over  and  through  a  per- 
forated copper  plate  placed  at  the  head  of  a  launder.  The  over- 
size of  the  screen,  made  up  of  leaf-copper  and  small  pieces  of 


OF   THE    COMMON    METALS.  283 

iron,  is  thrown  into  a  heap  and  picked  over  by  girls,  who  remove 
the  scrap-iron.  The  fine,  which  passes  through  the  screen,  is 
turned  over  against  a  stream  of  water  which  washes  out  the  dirt 
and  lighter  particles,  leaving  the  copper  behind.  For  a  few  yards 
at  the  head  of  the  launder  is  to  be  found  No.  i  precipitate  of  94% 
Cu,  0.3%  As,  then  comes  No.  2  precipitate  of  92%  Cu,  followed 
by  No.  3  precipitate,  quite  fine  and  containing  50%  Cu,  5%  As, 
the  graphite  of  the  pig  iron,  and  the  bulk  of  the  bismuth  and 
antimony  coming  from  the  liquors.  No.  I  and  2  precipitate  are 
sacked  for  shipment  and  No.  3  precipitate  is  added  to  a  blast-fur- 
nace matting-charge. 

The  old  Hunt-and-Douglas  process  for  the  extraction  of  copper. 

This  method  depends  upon  the  fact  that  copper  oxide  is  decom- 
posed by  ferrous  chloride  solutions,  forming  insoluble  ferric 
oxide,  while  the  copper  goes  into  solution  as  cuprous  and  cupric 
chlorides.  The  copper  is  then  precipitated  by  iron,  the  ferrous 
chloride  solution  being  regenerated,  and  needing  only  a  little  salt 
to  fit  it  for  use  again.  As  the  copper  must  be  in  oxidized  form  to 
go  into  solution,  the  ore  must  be  crushed  and  then  roasted  to  con- 
vert most  of  the  copper  into  oxide  or  sulphate. 

The  ore  is  dry-crushed  through  rolls,  to  about  4-mesh,  fine 
enough  for  roasting.  It  may  then  be  roasted,  preferably  in  one 
of  the  automatic  roasters  already  described.  The  roasted  ore  is 
placed  in  tanks,  which  have  been  carefully  coated  with  asphalt 
paint  to  resist  the  action  of  the  solutions.  The  percolating  solu- 
tion is  made  by  dissolving  120  parts  of  ferrous  chloride  in  1,000 
parts  of  water  and  adding  280  parts  of  green  vitriol  (FeSO4, 
6H2O).  To  this  solution  common  salt  is  added,  and,  by  stand- 
ing, sodium  sulphate  crystallizes  out,  leaving  a  mother  liquor, 
which  is  used  for  the  percolation.  The  reaction  by  which  the  cop- 
per oxides  in  the  ore  go  into  solution  is  as  follows : 

1.  3  CuO  +  2  FeCU  =  Cu,Cl2  +  CuCl2  +  Fe2Os.     While  the 
Cu2Cl2  is  not  soluble  in  water,  it  is  taken  up  in  presence  of  the 
common  salt  in  the  liquor. 

2.  3  Cu2O  +  2  FeCl2  =  2  Cu,Cl2  +  Fe2O3  +  2  Cu  and 

3.  2  Cu  +  2  CuCl2  =  2  Cu2Cl2,  which  again  is  soluble  in  the 
salt  liquor. 


284  THE    METALLURGY 

The  copper-containing  solution  passes  to  precipitating  tanks  or 
launders  containing  scrap-iron  or  pig  iron,  and  the  copper  is  pre- 
cipitated. The  consumption  of  iron  by  this  method  is  small  as 
compared  with  the  amount  needed  in  precipitating  copper  sul- 
phate, since  much  of  the  copper  is  present  as  cuprous  chloride. 
In  precipitating  copper  sulphate  2.5  to  3.5  Ib.  iron  is  needed  to 
precipitate  a  pound  of  copper. 

AVii'  Hunt-and-Douglas  process  for  the  extraction  of  copper. 

Referring  to  the  old  Hunt-and-Douglas  process  we  find  that 
the  copper  has  been  removed  from  the  roasted  ore  in  a  specially 
prepared  ferrous  chloride  solution.  In  the  new  process  instead  of 
using  scrap-iron  for  precipitation,  the  copper  is  obtained  by  pass- 
ing a  current  of  sulphurous  acid  gas  through  the  solution.  A 
white  heavy  cuprous  chloride  is  precipitated  which  can  be  readily 
decanted  and  washed.  Sulphuric  and  hydrochloric  acids  are  gen- 
erated in  the  solution,  which  only  needs  the  addition  of  some  com- 
mon salt  to  make  them  ready  for  further  use.  One  great  advan- 
tage of  the  method  consists  in  the  rapid  dissolving  of  the  oxidized 
copper  present  by  the  strongly  acid  solution,  which  attacks  even 
sulphides  with  considerable  energy.  And  lead  and  silver  present 
remain  undissolved. 

Neill  process  for  the  extraction  of  copper. 

The  process  depends  upon  the  use  of  sulphurous  acid  for  the 
leaching  out  of  the  copper.  It  is  preferably  used  on  oxidized 
ores  such  as  native  carbonates  and  oxides,  which  are  readily 
soluble  in  sulphurous  acid  with  the  formation  of  cuprous  sul- 
phite Cu2SO3.  The  salt  is  soluble  in  water  containing  sulphurous 
acid,  but  not  in  water  alone.  The  process  is  suitable  also  for  sul- 
phide ores  which  have  been  roasted.  Lime  and  magnesia  are  ob- 
jectionable because  they  use  up  some  of  the  sulphurous  acid. 

For  the  oxidized  ore,  crushing  is  performed  by  means  of  rolls, 
which  reduce  it  to  2o-mesh  size.  It  is  then  charged  into  leaching 
tanks,  and  is  mixed  with  water  containing  SO2  gas.  The  sulphur- 
ous acid  is  generated,  preferably  by  roasting  sulphide  ores  in  a 
pyrite-roaster,  together  with  a  sufficient  supply  of  air.  The  gases 
thus  produced  and  containing  3  to  5%  SO2,  are  driven  through  the 
pulp  in  the  tank  by  means  of  an  air-compresser,  thoroughly  satu- 


OF   THE    COMMON    METALS.  285 

rating  the  whole.  From  the  tank  the  ore  passes  to  filter-presses 
where  the  solution  is  removed,  the  residual  tailing  being  then  sent 
to  waste.  The  solution  passes  now  to  precipitating  tanks,  where 
it  is  heated  by  steam  to  the  boiling  point,  thus  driving  off  the  SO2 
gas.  In  consequence,  the  copper  is  precipiated  as  cupro-cupric 
sulphite  (CuSO3,  Cu2SO3+H2O)  a  heavy  crystalline  compound 
of  a  dark  red  color  and  containing  49.1%  Cu.  The  solution  from 
the  precipitating  tank  is  now  run  to  vats  containing  scrap-iron, 
as  a  precaution  to  ensure  the  removal  of  the  last  traces  of  copper. 
The  precipitate  readily  settles  out  of  the  solution,  is  washed  by 
decantation,  dried  and  reduced  and  melted  to  metallic  copper  in 
a  reverberatory  furnace.  The  process  has  the  advantage  that  a 
unit  of  copper  converted  into  cuprous  sulphide  needs  but  one-half 
the  sulphur  required  to  convert  it  into  cuprous  sulphate.  Cuprous 
sulphite  is  here  precipitated  from  its  solution  without  the  use  of 
scrap-iron,  a  great  advantage  in  remote  districts  where  transpor- 
tation would  be  high.  Sulphurous  acid  acts  but  little  on  other 
metals,  and  thus  furnishes  a  purer  copper. 

83.     THE  COPPER  CONVERTER. 

In  the  making  of  steel  by  the  bessemer  process  the  upright  con- 
verter is  used,  but  in  performing  the  same  operation  upon  copper, 
the  trough  or  barrel  converter  is  preferred,  called  also  the  Leg- 
horn converter,  from  the  city  in  Italy  where  it  was  first  used. 

Fig.  121  represents  such  a  converter  with  the  mechanism  for 
revolving  or  tilting  it.  It  is  made  of  heavy  steel-plate  in 
two  sections,  so  that  the  top  can  be  lifted  off  leaving  the 
interior  accessible  for  relining.  To  one  side  of  the  shell 
is  bolted  a  cast-iron  wind-box  fitted  with  tuyeres.  To 
one  head,  concentric  with  the  axis,  is  bolted  a  cast-steel 
open  gear,  which  engages  with  a  steel  rack  connected  to 
the  piston  of  a  hydraulic  cylinder  carrying  a  pressure  of 
300  Ib.  to  the  square  inch,  and,  by  means  of  this  rack,  the 
converter  is  tilted.  'Carrying  rings'  made  of  heavy  T-rail  are 
bolted  to  the  shell  and  rest  upon  four  friction-rollers  carried  by 
heavy  cast-iron  base-frames.  Two  inclined  tie  links  will  be 
noticed  at  the  front  of  Fig.  121  which  hold  the  rack  meshed  into 
the  gear.  When  the  converter  is  to  be  removed  it  is  lifted  off  its 


286  THE    METALLURGY 

rollers  by  a  4O-ton  overhead  traveling-crane.  The  rack,  which  is 
jointed  at  the  head  of  the  piston,  is  then  swung  forward  into  gear 
by  turning  an  eccentric  pin  at  the  foot  of  the  tie-links.  A  four- 
way  valve,  adjoining  the  cylinder  and  controlled  by  a  hand  wheel 
(seen  next  to  the  rack,  Fig.  121),  serves  to  operate  the  converter. 
There  are  four  eyes  riveted  both  to  the  converter  and  to  its  top 
by  which  the  four-fold  lifting  chain  is  attached. 

This  illustration  shows  a  two-stand  converter-plant.  In  the  fore- 
ground, at  the  left,  are  the  matte  ladles  which  are  handled  by  the 
crane.  Above  and  back  of  the  converters  is  the  outlet  flue  where 
the  fumes  from  the  converters  escape.  Hoods  attached  to  the 
front  side  of  the  flue  receive  the  fumes. 

84.     CONVERTING  COPPER  MATTE  TO  BLISTER  COPPER. 

Before  a  converter  receives  its  first  charge  of  molten  matte,  it 
has  been  dried  out  and  heated  to  a  low  red  by  a  fire  of  wood,  fol- 
lowed by  coke,  urged  by  a  blast  introduced  at  the  tuyeres.  A 
newly  lined  converter  of  the  trough  type,  7  ft.  in  diam.  by  10  ft. 
6  in.  long,  will  take  as  its  initial  double-charge  5  tons,  and  a  final 
double-charge,  just  before  relining,  of  12  tons.  A  charge  for  a 
converter  is  tapped  from  a  furnace  forehearth  or  from  a  rever- 
beratory  furnace  into  a  steel  ladle  managed  by  an  electric  travel- 
ing-crane. The  charge  is  then  brought  to  the  converter  and 
poured  into  its  spout.  (The  ladles  are  5  ft.  diam.  by  3  ft.  10  in. 
high,  are  lined  with  ordinary  loam  plastered  on  by  hand  and  dried 
by  a  wood  fire.)  The  converter  is  turned  into  its  normal  posi- 
tion, a  full  pressure  of  blast  being,  at  the  same  time,  let  on. 

Now  begins  the  first,  or  slag-forming,  stage  of  the  process. 
The  end  of  this  stage  is  determined  by  the  appearance  of  the  is- 
suing flame,  the  greenish  border  of  the  flame  having  given  place 
to  a  pale,  permanent  blue  as  all  the  iron  becomes  oxidized.  Sup- 
posing that  we  are  treating  a  50%  matte  it  would  be  in  composi- 
tion much  as  follows,  50*7^  Cu,  20%  Fe  and  24%  S,  and  corre- 
sponding to  the  formula  or  Cu2S,  FeS,  together  with  a  little 
Fe3O4.  As  the  blowing  proceeds  the  following  reactions  occur. 
The  FeS  is  changed  to  FeO  as  follows: 

FeS  +  3  O  =  FeO  +  SO2 
23800  66400       71000=113400 


OF    THE    COMMON    METALS. 


287 


FIG.  121.     COPPER  CONVERTER. 


288  THE    METALLURGY 

This  FeO  takes  up  silica  from  the  lining  thus : 
FeO  +  SiO2  =  FeSiO3, 

and  forms  a  slag,  into  which  also  goes  the  alumina  of  the  lining. 
The  slag,  thus  formed,  is  poured  from  the  converter  into  a  ladle 
by  turning  it  down.  As  the  slag  flows  out,  the  converter  man 
passes  a  rabble  through  it  from  time  to  time.  He  can  tell  when 
the  matte  is  beginning  to  escape,  whereupon  he  signals  for  the 
converter  to  be  returned  to  its  normal  position,  and  the  blast  pres- 
sure to  be  turned  on,  in  order  to  continue  the  blow.  At  this  stage 
the  copper  has  been  brought  up  to  white  metal  of  76%  copper. 
The  slag,  still  containing  1.5  to  2.%  of  copper,  and,  say,  0.5  to  i.o 
oz.  silver  per  ton,  is  sent  back  to  the  furnaces  to  be  re-treated  to 
recover  the  copper  and  silver.  Some  pieces  of  cold  matte  may  be 
thrown  into  the  converter  to  help,  by  oxidation,  to  make  the 
charge  hotter,  and  with  the  slag  not  wholly  poured  out,  to  make 
a  covering  for  the  metal.  Under  normal  conditions  the  escaping 
flame  is  of  a  bluish  white,  gradually  changing  to  a  rose  red,  and 
finally  to  a  brownish  red.  It  lessens  in  length  and  volume  until, 
at  last,  there  is  only  a  brick  red  flicker  in  the  escaping  gases.  The 
determination  of  the  completion  of  the  operation  requires  care 
and  experience,  since,  if  carried  too  far,  over-blown  copper  may 
result.  In  this  second  stage  of  blowing,  the  oxidation  of  the  sul- 
phur still  retained  by  the  copper,  proceeds,  so  that  we  have 
Cu,S  +  30  =  Cu,O  +  SO2,  and  the  Cu,O  thus  formed,  reacting 
on  other  Cu2S,  produces  copper  with  an  evolution  of  sulphur  di- 
oxide as  follows :  Cu,S  -+-  2  Cu2O  =  4  Cu  +  SO2.  So  far  as  the 
agitation  by  the  air  will  permit  it,  the  molten  material  in  the  con- 
verter tends  to  separate  into  layers  as  follows:  An  increasing 
layer,  of  copper  at  the  bottom,  a  decreasing  layer  of  matte  above 
it,  and  a  cover  of  slag.  The  blast  entering  at  the  side  tends  to 
blow  through  the  matte  floating  above  the  forming  copper. 

So  soon  as  the  contents  of  the  converter  have  been  changed  to 
copper,  the  converter  is  turned  down,  the  blast  shut  off,  and  the 
copper  poured  into  molds,  which  have  been  mounted  on  a  carriage 
running  on  a  track  beneath  the  converter.  The  resulting  pigs 
weigh  250  to  300  Ib.  each.  The  converter  is  now  turned  back  into 
the  receiving  position  and  the  next  charge  promptly  poured  in. 


OF   THE    COMMON    METALS.  289 

The  first  charge  for  a  newly  lined  converter  is  necessarily 
small,  but,  as  the  eating  out  of  the  lining  proceeds,  and  more  room 
is  given,  it  is  greatly  increased,  as  has  already  been  stated. 

In  converting  the  higher  grade  matte,  the  process  as  above  out- 
lined is  simple,  but  for  lower  grade  mattes  of  35  to  45%  copper, 
the  methods  are  modified  by  'doubling,'  as  it  is  called.  Thus,  in  a 
newly  lined  converter,  a  charge  of  2  tons  is  introduced,  and  this 
is  blown,  through  the  first  stage,  to  white  metal.  The  converter 
is  turned  down,  the  slag  poured  off  and,  upon  its  return  to  the  re- 
ceiving position,  3  tons  more  of  matte  is  added,  making  a  double 
charge  of  5  tons  in  all.  This  is  blown,  the  slag  removed  and  then 
finished  to  blister  copper.  The  advantage  of  this  procedure  lies 
in  the  fact  that  there  is  an  increased  output,  and  that  there  is 
enough  white  metal  produced  and  blister  copper  formed  to  prop- 
erly fill  the  converter  and  to  make  a  sufficient  pouring  of  blister 
copper. 

The  final  charge  is  given  at  12  tons,  and  since  large  charges  are 
the  most  profitable,  there  is  a  temptation  to  press  operations  to  the 
limit.  The  lining  becomes  dangerously  thin,  showing  a  red  hot 
spot  on  the  shell.  Sometimes  this  spot  can  be  kept  cool  by  a 
stream  of  water  from  a  hose,  and  the  charge  finished,  but  if 
not,  the  charge  must  be  poured  out  and  transferred  to  another 
converter. 

The  time  taken  for  converting  copper  matte  will  average  40  to 
45  minutes  for  the  first  and  50  minutes  for  the  second  blow;  20 
to  30  minutes  are  needed  for  the  charging,  skimming,  and  poling, 
or  we  may  say  that  a  blow  will  last  2  hours  in  all.  On  a  new 
charge,  it  may  be  as  little  as  an  hour,  and  for  a  final  charge,  as 
high  as  two  hours.  A  converter  will  last  from  5  to  9  charges, 
according  to  the  grade  of  matte  and  the  durability  of  the  lining. 
Low-grade  matte,  of  course,  corrodes  the  lining  most  rapidly. 
The  blast  pressure  may  average  13  Ib.  per  square  inch. 

85.     RE-LINING  THE  CONVERTER  VESSELS  OR  SHELLS. 

The  old  way  of  cooling,  after  removing  the  top,  was  to  cool 
the  interior  of  the  vessel  by  means  of  a  stream  of  water  from  a 
hose,  and  then  re-line.  The  objection  to  this  is  that  the  water, 


29O  THE    METALLURGY 

taking  up  copper  salts,  would  have  a  corrosive  action  on  the  shell, 
eventually  weakening  it;  consequently  it  has  been  abandoned. 
One  may  see  the  effect  of  such  corrosive  action  upon  the  rails  be- 
neath the  converter  where  they  have  been  frequently  wet  from 
spilled  or  escaping  water. 

The  lower  section,  being  set  on  its  side  by  means  of  the  crane, 
is  cleaned  from  adhering  matte  and  slag  by  cutter  bars  so  that 
the  new  lining  may  have  a  new  and  clean  surface  of  attachment. 
The  body  is  then  set  upright  and  the  bottom  lining  tamped  in  to 
within  6  in.  of  the  tuyeres.  A  sectional  rectangular  tapering  iron 
shell  or  form  is  now  put  in  to  make  the  interior  cavity.  Around 
this  the  lining  mixture  or  ganister  is  firmly  tamped  by  means  of 
heated  iron  tampers  or  rammers  to  the  top  of  the  form.  The 
form  is  removed  and  the  tuyere  openings  are  now  punched 
through.  The  top  section  is  put  on  and  bolted  fast.  The  work- 
man then  enters  the  converter  and  completes  the  lining  with  balls 
of  ganister  composed  of  72%  quartz  and  28%  clay.  The  shell  is 
dried  out  by  putting  in  a  fire  of  a  little  wood,  followed  by  coke, 
a  light  fan  blast  being  also  admitted  at  the  tuyere  openings.  It  is 
important  to  have  enough  converters,  that  they  may  be  thor- 
oughly dried  out  and  heated  before  using,  since,  otherwise,  the 
life  of  the  lining  is  much  lessened.  The  shell  is  now  ready  to  place 
upon  the  stand. 

Re-melting  copper  matte. 

The  older  way  of  preparing  copper  matte  for  converting  was  to 
re-melt  the  matte,  which  had  been  tapped  off  into  molds  at  the 
reverberatory  or  blast-furnace.  This  was  done  in  a  circular 
blast-furnace  or  cupola,  such  as  is  used  for  melting  pig  iron  at  the 
iron  foundry.  The  furnace  was  driven  at  a  rate  such  as  would 
give  the  needed  supply  of  matte  to  the  converters,  a  somewhat 
deep  crucible  being  used.  The  matte  was  tapped  in  sufficient  vol- 
ume to  fill  the  converter,  and  if  there  was  not  enough  to  do  this 
at  one  tapping,  more  was  drawn  off  as  fast  as  it  could  be  melted. 
Meanwhile  the  converter  would  be  blowing  the  first  portion.  The 
matte  flowed  from  the  cupola  by  means  of  a  clay-lined  launder  to 
the  converter. 


OF   THE    COMMON    METALS. 


291 


THE    METALLURGY 

Necessity  of  a  silicious  lining. 

For  the  reason  that  FeO  is  developed  as  the  result  of  the  blow, 
it  must  have  silica  with  which  to  combine.  Were  it  not  for  its 
presence,  an  infusible  mass  of  iron  oxide  would  soon  accumulate 
within  the  converter.  It  would  not  be  possible  to  run  writh  a  basic- 
lined  or  water-jacketed  shell.  This  has  been  several  times  under- 
taken, since  it  would  save  the  very  considerable  expense  and  time 
lost  in  re-lining,  but  such  experiments  have  always  failed.  The 
life  of  the  lining  is  less  as  the  grade  of  matte  is  lower  because, 
with  so  much  FeO  coming  from  it,  a  more  basic  slag  is  formed. 
The  lining  also  falls  off  in  lumps  owing  to  this  higher  corrosive 
action.  When  converting  60%  Cu  matte,  the  slag  will  contain 
40%  SiO2,  while  with  35%  Cu  matte,  the  resultant  slag  will  be 
found  to  contain  but  25%  SiO2. 

'  86. .   ACCESSORIES  OF  A  CONVERTER  PLANT. 

The  size  of  a  converter  plant  is  indicated  by  the  num- 
ber of  stands  or  stalls  at  which  copper  can  be  blown. 
Fig.  123  and  124  are  illustrations  of  a  two-stand  plant. 
In  addition,  there  should  be  at  least  two  extra  shells  for  each 
stand.  Thus,  while  one  shell  is  in  place  and  being  used,  the  two 
others  are  being  taken  apart,  cleaned  out,  re-lined  and  dried, 
ready  to  take  the  place  of  an  exhausted  shell  from  the  stand. 
To  handle  these  shells,  to  bring  matte  to  the  converters,  and  to 
take  away  slag  (and  copper  also  where  a  re-melting  furnace  is 
used)  the  steady  service  of  a  4O-ton  electric  crane  is  needed.  This 
crane  (see  Fig.  123  and  124)  having  a  span  of  40  ft.  travels  from 
end  to  end  of  the  converter-house.  Both  matte  and  slag  are 
transferred  from  the  forehearth  of  the  blast-furnace,  or  from  the 
reverberatory,  by  means  of  cast-steel  ladles  (Fig*.  125),  which 
have  been  lined  with  a  clayey  loam.  The  ladle  should  be  large 
enough  to  take  a  charge  of  5,000  to  12,000  Ib.  of  the  matte  which 
is  to  be  poured  into  the  converter.  The  ganister  for  lining  may 
be  a  mixture  of  quartz  rock,  coarsely  ground  in  a  Chilean  mill  or 
wet  pan  (Fig.  126),  together  with  15^  of  clay,  or  it  may  consist 
of  a  silicious  gold  or  silver-bearing  quartz  ore  mixed  with  clay. 
The  advantage  of  the  use  of  the  ore  is,  that  its  contained  values  in 


OF   THE    COMMON    METALS. 


293 


294 


THE    METALLURGY 


gold  or  silver  go  into  the  copper  as  the  lining  is  consumed.  To 
blow  the  air  at  a  pressure  of  8  to  15  Ib.  per  square  inch,  a  horizon- 
tal blowing  engine  is  generally  used,  capable  of  delivering  300 
cu.  ft.  of  free  air  per  tuyere  of  the  converter  or  for  a  converter 
as  shown  in  Fig.  127,  2,500  to  3,000  cu.  ft.  of  air  per  minute. 
Beside  the  Chilean  mill  (or  wet  pan),  already  mentioned,  there 
is  a  rock-breaker  for  coarsely  crushing  the  quartz  rock,  and  a 
fan  for  supplying  air  under  low  pressure  to  the  fires  in  the  con- 
verter shells,  which  have  been  lined  and  are  being  dried  out.  The 
converters  are  operated  by  a  hydraulic  cylinder  supplied  by  water 


FIG.  125.     CAST-STEEL  LADLE. 

under  a  pressure  of  175  to  500  Ib.  per  square  inch.  To  regulate 
this  pressure  a  hydraulic  accumulator  (Fig.  128)  is  employed. 
This  consists  of  a  fixed  vertical  cylinder  in  which  slides  a  weighted 
plunger.  When  the  water  is  quickly  used  the  plunger  falls,  thus 
regulating  the  supply  according  to  needs  and  keeping  up  the  re- 
quired pressure.  The  water  is  supplied  by  a  duplex  pump,  as 
shown  in  the  figure,  pumping  against  the  weighted  plunger. 

Where  it  has  been  attempted  to  cast  copper  from  the  converters 
into  molds  for  the  formation  of  anodes,  the  molten  metal  is  in 
such  an  agitated  condition,  owing  to  the  occluded  gases,  that  the 


OF   THE    COMMON    METALS. 


295 


296  THE    METALLURGY 

anodes  are  quite  rough,  and  liable,  through  irregularities  both  in 
form  and  in  composition,  to  tend  to  short  circuiting  and  to  poor 
working  in  the  electrolytic  bath.  To  obtain  smoother  anodes  a 
reverberatory  furnace  has  been  used.  Into  this  reverberatory  the 
converter  copper  is  transferred  as  it  is  made,  and  every  morning 
the  accumulated  copper  is  poled  to  reduce  the  contained  copper 
oxide,  and  molded  into  anodes  as  in  ordinary  copper  refining. 
While  this  operation  is  going  on,  of  course  no  copper  can  be  put 
in  the  reverberatory.  Consequently  it  is  poured  into  ingot  molds, 
as  in  the  older  way  of  casting.  \Yhen  the  anode  casting  is  com- 
pleted, and  the  furnace  empty,  the  ingots,  thus  just  cast,  are  added 
in  the  reverberatory  to  the  next  charge. 

87.     LOSSES  ix  BESSEMERIZIXG  OR  CONVERTING  COPPER  MATTE. 

The  fumes  from  the  converter  consist  mainly  of  nitrogen,  sul- 
phurous acid  and  traces  of  volatilized  metals  (As,  Sb,  Te,  Pb, 
Zn,  also  Cu  and  Ag).  The  loss  of  gold,  as  in  pyritic  smelting  is, 
however,  exceedingly  small,  while  the  loss  of  silver  depends  upon 
the  amount  of  volatile  metals  present,  as  just  given.  A  portion 
of  the  values  in  copper  and  silver  are  recovered  in  the  flues.  Thus 
the  silver  in  the  flue-dust  was  found  to  be  as  follows :  At  the 
branch  from  the  converter  hood,  28  to  64  oz. ;  at  the  first  part  of 
flue-chambers,  22  to  65  oz. ;  further  along  the  flue,  19  to  46  oz. ; 
and  near  the  stack,  17.5  oz.  per  ton.  Thus  the  larger  the  chamber 
the  better  the  recovery  of  values.  The  losses  in  treatment  as  the 
result  of  extended  runs  amount  to  I  to  1.5%  of  the  copper  and 
2  to  2.5%  of  the  silver  treated.  In  attempting  to  treat  concen- 
tration matte  coming  from  lead-silver  blast-furnaces,  and  con- 
taining 40%  Cu  and  10.2%  Pb,  the  silver  losses  were  very  serious, 
amounting  to  33  to  40% ,  due  to  the  volatilization  loss  of  the  lead, 
which  would  be  all  practically  driven  off,  as  is  the  case  when  lead 
is  present  in  a  copper-matting  charge  of  a  blast-furnace. 

88.     COST  OF  CONVERTING. 

Hixon,  in  his  'Notes  on  Lead  and  Copper  Smelting  and  Refin- 
ing/ gives  the  following  as  the  cost  of  converting  matte  per  pound 
of  copper : 


OF    THE    COMMON    METALS. 


297 


298  THE     METALLURGY 

Re-melting   matte    o.2oc 

Labor  and  re-lining  converters 0.25 

Labor  on  converter    o.io 

Re-smelting  converter  slag   0.05 

Supplies    0.05 


or  for  a  plant,  using  direct  matte,  0.45  to  O.4OC. 

Fig.  127  is  a  duplex  compound  blowing  engine  for  supplying 
air.  Such  an  engine  would  have  steam  cylinders  17  and  32  in. 
diam.  respectively,  and  air  cylinders  36  in.  diam.,  all  having  a 
stroke  of  30  in.  and  delivering  5,600  cu.  ft.  of  air  per  minute  at 
a  pressuie  of  10  Ib.  per  sq.  in.  and  be  capable  of  supplying  the 
converter  plant  (Fig.  123). 

89.     A  3oo-Tox  COPPER  SMELTING  AND  REFINING  PLANT. 

Fig.  123  is  a  sectional  elevation,  and  Fig.  124  a  plan  of  a  plant 
(already  referred  to)  where  copper-bearing  ore  is  treated  in  two 
furnaces,  each  42  by  144  in.  section  at  the  tuyeres,  and  the  matte, 
as  fast  as  it  accumulates,  is  transferred  to  one  or  other  of  the 
converters  in  the  converter  building. 

Referring  to  Fig.  123  the  ore,  coming  in  by  car  on  an  elevated 
track,  discharges  into  inclined-bottom  bins.  Another  overhead 
track  at  the  right  receives  the  coke.  These  materials  are  drawn 
off,  weighed  and  discharged  to  the  blast-furnaces.  The  furnace- 
top  has  two  outlets,  one  discharging  directly  to  the  outside  above 
the  roof,  the  other  entering  a  horizontal  flue  5  ft.  wide  by  7  ft. 
deep  where  the  flue-dust,  as  it  accumulates,  may  be  drawn  off. 
This  flue  crosses  the  roof,  and  passing  under  the  coke-track,  ter- 
minates in  a  high  sheet-iron  brick-lined  stack.  The  settlers  of  the 
blast-furnace,  10  ft.  diam.  by  5  ft.  high  (shown  also  in  Fig.  124) 
receive  the  flow  of  the  blast-furnaces,  the  slag  being  then  taken 
from  the  settler  into  a  slag  car,  which  comes  by  track  close  to  the 
settler,  while  the  matte  is  tapped  from  its  lower  part  to  a  matte- 
ladle  handled  by  the  crane. 


OF   THE    COMMON    METALS. 


299 


FIG.  128.    HYDRAULIC  ACCUMULATOR. 


PART  VII.    LEAD 


PART   VII.     LEAD. 


90.     THE  ORES  OF  LEAD. 

Lead  ores  are  those  in  which  -lead  is  the  principal  constituent, 
though  the  term  is  also  applied  to  those  mineral  aggregates  in 
which  there  is  over  10%  lead.  The  ores  of  lead  may  be  divided 
into  two  classes,  sulphide  ores  and  oxidized  ores,  but  it  must  be 
understood  as  a  term  applied  to  the  constituent  which  is  in  ex- 
cess, since,  in  many  lead  ores,  both  sulphides  and  oxides  are  to  be 
found  mingled.  Ores  containing  no  lead  are  called  dry,  and  those 
with  lead,  leady.  This  latter  term  is  the  converse  of  dry,  so 
that  one  would  never  call  a  leady  ore  a  'wet'  one. 

Galena,  when  pure,  contains  86.6%  Pb,  13.4%  S.  It  comes, 
mixed  with  more  or  less  vein  matter  or  gangue,  and,  in  prepara- 
tion for  the  smelter,  this  may  be  largely  removed  by  ore-dressing 
methods.  The  table  herewith  gives  an  idea  of  such  ores,  both  be- 
fore and  after  dressing. 

Galena  Ores. 


Raw  Ore 

Concentrate 

Pb 

Pb 

Ag 

< 

* 

Oz.  per  ton 

Rockville,  Wis. 

0.3 

St.  Joseph,  Mo. 

7.0 

70.0 

Kellog-g,  Idaho. 

11.0 

60.0 

30.0 

Kansas  and  Missouri  galena  contains  but  little  silver,  but  in 
the  Rocky  Mountain  region  it  is  not  only  argentiferous,  but  con- 
tains a  little  gold,  the  precious  metals  determining  its  value  as 
much  as  the  lead.  Metallic  sulphides  are  often  associated  with 


304 


THE    METALLURGY 


galena  and,  (with  the  gangue,)  may  carry  precious-metal  values. 
This  may  interfere  with  concentrating  the  ore,  and,  hence  that 
operation  may  have  to  be  omitted,  but  an  ore  of  35  to  40%  Pb  is 
a  very  acceptable  one  at  the  smelter.  If  of  this  tenor  in  lead, 
and  with  no  associated  sulphides,  it  would  contain  no  more  than 
5%-  sulphur,  so  little,  in  fact,  as  to  warrant  its  direct  treatment 
in  the  blast-furnace  without  preliminary  roasting. 

Oxidized,  also  called  carbonate  ores.  They  include  anglesite 
PbSO4,  containing,  when  pure,  73.6%  Pb,  and  cerussite,  PbCO3, 
containing  also,  when  pure,  83.5%  Pb.  When  in  sandy  or  earthy 
form,  these  are  called  sand  or  soft  lead  carbonates,  but  when  in 
hard  lumps  or  in  stony  form,  they  are  known  as  hard  lead  car- 
bonates. Lumps  of  ore,  originally  galena,  may  be  found  more 
or  less  profoundly  altered  into  anglesite  or  cerussite.  The  fol- 
lowing table  gives  the  constitution  of  various  carbonates.  Silver 
occurs  in  oxidized  ores  in  the  form  of  chloride,  the  gold  probably 
in  the  native  state.  There  are  various  other  lead  minerals,  but  so 
small  in  quantity  as  not  to  constitute  ores. 

Carbonate  Ores. 


Pb 

so, 

* 

Fe 

* 

CaO              S 

*                  * 

Ag 

Oz. 

per  ton 

Southwest  Missouri 

72.0 

Leadville,  Colo. 

38.0 

25.0 

Leadville,  Colo. 

21.0 

65.0 

Red  Mountain,  Colo. 
Eureka,  Xev. 

17.0 
33.2 

3.0 

24.1 

1.1 

128.0 
128.0 

91.     BEDDING  ORES. 

In  a  custom-works  ores  of  many  kinds  are  received,  often  in 
small  lots.  Such  ores  are  'bedded'  in  large  bins,  holding  several 
hundred  tons  each,  and  when  so  bedded  are  treated  as  one  ore. 
Each  kind  of  ore  is  unloaded  into  the  bin  and  is  then  spread  out 
in  an  even  layer  upon  its  predecessor,  which  has  been  treated  in 


OF    THE    COMMON    METALS.  305 

the  same  manner.  The  ore,  when  used,  is  shoveled  up  from  the 
floor,  the  upper  portion  falling  down  and  mixing  with  the  lower 
layers.  Thus  a  uniform  mixture  of  all  the  ore  of  the  bed  is 
obtained  for  use  at  the  furnace.  The  advantages  are  that  a  charge 
can  be  calculated  and  adjusted,  after  which  it  remains  the  same, 
often  for  several  days,  and  again,  the  values  are  as  one  ore  upon 
the  books  of  the  company. 

In  the  laboratory  the  aggregate  analysis  is  made  as  follows :  A 
list  of  ores  and  of  their  dry  weights  is  prepared.  The  chemist 
weighs  out  on  his  balance  from  the  reserved  samples  of  each  of 
the  ores  an  amount  proportionate  to  the  quantity  of  that  ore  in 
the  bin.  The  total,  amounting  to  one  or  two  ounces,  is  thoroughly 
mixed,  and  from  it  the  portions  are  taken  for  the  determination 
of  SiO2,  Fe,  CaO  and  S. 

It  may  also  be  determined  for  Ag,  Au  or  Pb,  though  these 
quantities  are  more  often  estimated  from  the  assay  of  the  indi- 
vidual ores.  The  determinations  thus  made  are  used  in  calculat- 
ing the  charge.  Besides  ores  bedded  in  this  way,  lots  which  would 
fill  a  large  bin  remain  unmixed  and  are  treated  as  a  separate  item 
of  the  charge.  The  same  is  true  of  smaller  lots  of  which  a  mode- 
rate amount  is  taken  for  each  charge,  and  these  are  called  'side 
ores.'  Roasting  ores  are  also  preferably  bedded  because,  when 
uniform,  they  are  better  known  and  handled  in  the  roaster,  and 
are  more  accurately  smelted.  Such  beds  may  be  prepared  to  con- 
tain the  proper  amounts  of  pyrite,  silica  and  lead  which  work  best 
when  roasted.  . 

Crushing  and  bedding  of  lead-bearing  ores  for  roasting. — 
This  is  often  performed  in  two  stages.  The  coarse  crushing  of 
lumpy  ore  is  best  done  through  ore-breakers,  either  of  the  jaw 
or  gyratory  type,  the  ore  being  reduced  to  0.75  in.  It  is  then 
passed  through  rolls  36  in.  diam.  by  14  in.  face,  where  it  is 
crushed  to  pass  a  No.  3  to  a  No.  10  mesh  screen,  according  to  its 
nature.  A  pyritiferous  ore  need  be  crushed  no  finer  than  No.  3 
mesh,  many  ores  and  matte  to  No.  5,  while  galenas  and  zinciferous 
ores  need  finer  crushing  to  No.  10  mesh. 

Ores  are  not  to  be  roasted  indiscriminately,  but  are  preferably 
bedded  or  placed  in  a  large  bin  in  layers  so  that  the  roasting  may 
be  the  same  for  a  considerable  time,  in  order  that  the  men  may 


306 


THE    METALLURGY 


know  how  to  roast  it  to  the  best  advantage.  Also,  by  so  doing, 
ores  of  different  constitutions  may  be  combined.  For  instance,  a 
leady  ore  is  better  roasted  mixed  with  a  silicious  sulphide,  pyrite 
assists  a  leady  or  zincky  ore,  zinc  makes  the  mixture  more  infu- 
sible, so  that  it  will  not  so  soon  soften  at  the  finish  of  the  roast.  A 
bed  formed  to  contain  10  to  15%  SiO2,  20  to  28%  Fe  and  20  to 
28%  Pb  works  quite  well.  Mixtures,  containing  less  lead  than 
this,  roast  more  readily,  but  are  quite  pulverulent  when  roasted, 
and  make  much  flue-dust,  while  with  the  above  specified  propor- 
tion of  lead,  they  sinter  together  somewhat  and  thus  give  a  better 
product  for  the  blast-furnace. 


92.     REVERBERATOR  Y  LEAD  SMELTING. 

The  treatment  of  lead  ores  in  the  reverberatory  furnace  has 
not  made  headway  in  the  United  States  for  two  reasons.  In  the 
silver-lead  districts  the  ores  have  not  been  sufficiently  high  grade 


;E 


"iD^-i^^^H— *»•— 4   5 

HORIZONTAL  SECTION  ON  LINE  G,  H. 


FIG.    129.      LEAD-SMELTIXG    REVERBERATORY  FURNACE  (PLAN). 

in  lead  to  warrant  such  treatment,  and  lead  ore  has  been  much  in 
demand  as  a  collector  to  mix  with  other  ores.  Second,  in  the 
Mississippi  valley,  where  silver-free  high-grade  ores  occur,  the 


OF   THE    COMMON    METALS.  307 

question  of  skilled  labor  has  had  some  influence,  perhaps  un- 
justly so. 

The  roast  reaction  method. — High-grade,  pure,  non-silicious 
ores  are  best  treated  by  this  method.  Fig.  129  and  130  represent 
one  of  the  larger  and  more  recent  of  these  furnaces,  16  by  9  ft. 
hearth  dimension  and  having  a  fire-box  8  ft.  by  I  ft.  8  in.,  or  of 


Firebrick 


LONGITUDINAL  SECTION  ON  LINE  A,  B. 


FIG.  130.    LEAD-SMELTING  REVERBERATORY  FURNACE  (ELEVATION). 

14  sq.  ft.  area.  The  bottom  of  the  hearth  has  a  slope  from  the 
fire-bridge  to  the  external  well  or  basin  F,  located  at  the  cooler 
end  of  the  furnace.  The  flame  passes  to  the  stack  by  the  ports  at 
A.  The  charge  is  dropped  into  the  furnace  from  a  hopper  and 
through  a  hole  at  the  middle  of  the  roof.  There  are  four  work- 
ing-doors on  each  side,  where  the  charge  can  be  spread  out,  raked 
and  withdrawn.  The  lead  from  it  drains  to  the  basin,  F,  from 
which  it  is  dipped  out  from  time  to  time  as  it  accumulates,  and 
is  molded  into  ingots  or  bars. 

Operation. — The  working  of  the  furnace  may  be  divided  into 
two  stages,  namely,  oxidation  and  reduction. 

Oxidation. — Four  tons  of  ore,  crushed  to  5-mesh,  are  spread 
out  in  a  layer  3  in.  thick  upon  the  hearth,  and  heated  gradually 
to  just-visible  red  (500  to  600°  C).  The  roasting  which  takes 
three  to  four  hours  is  carried  on  until  partly  completed,  according 
to  the  reaction 

2  PbS  +  7  O  =  PbO  +  PbSO4  +  SO2 

The  temperature  is  kept  low,  and  the  charge  is  frequently 
raked  to  expose  new  surfaces  to  the  action  of  the  air  and  to 
prevent  agglomeration. 


308  THE    METALLURGY 

Reduction. — The  grate  is  filled  with  coal  to  give  a  neutral 
flame,  and  the  temperature  is  raised  to  700°  C  so  that  the  oxygen 
compounds  may  react  on  the  unchanged  galena  thus : 

PbS+2  PbO=3Pb+SO2 
PbS+PbSO4=2  Pb+2  SO2 

The  charge  gradually  softens,  but  not  to  melting,  white  fumes 
are  given  off  and  the  lead  begins  to  flow.  To  stiffen  the  charge, 
making  it  less  fusible  and  more  spongy,  slacked  lime  is  added 
and  stirred  in.  Rabbling  the  charge  is  also  performed  at  inter- 
vals to  promote  the  reaction.  By  this  time  the  flow  of  lead 
ceases,  but  the  pastry  residue  still  contains  about  half  of  the 
original  lead. 

To  extract  more  lead  a  second  roasting  takes  place,  followed 
by  a  second  re-acting.  It  takes  several  repetitions  of  the  process 
to  extract  the  bulk  of  the  lead.  Toward  the  end  there  will  be 
no  sulphide  left  to  re-act  on  the  oxides  and  sulphates.  To  reduce 
these,  coal  or  charcoal  is  mixed  in,  and  a  further  portion  of  lead 
obtained.  Each  successive  operation  will  be  shorter  than  the 
last,  and  the  temperature  will  be  carried  a  little  higher.  The 
lead,  as  it  flows  away  from  the  charge,  is  received  into  the  outer 
basin  F,  and,  after  skimming,  is  molded  into  bars  or  ingots.  The 
residue,  after  the  extraction  of  all  the  lead  possible,  is  a  gray  slag, 
still  containing  12  to  30%  Pb,  and  amounting  to  25%  of  the 
charge.  It  is  generally  sent  to  a  blast-furnace  where  the  lead  can 
be  thoroughly  removed. 

It  takes  upward  of  12  hours  to  work  through  a  charge  with  a 
consumption  of  45%  of  its  weight  of  coal. 

The  process  is  only  suited  to  concentrate  and  to  ores  rich  in 
lead,  not  containing  over  4  or  5%  SiO2,  which  ingredient  forms  a 
silicate  with  the  lead  oxides.  Small  amounts  of  the  metallic  sul- 
phides are  not  harmful;  indeed  pyrite  is  beneficial  at  the  first 
stage,  while  limestone,  dolomite,  blende  and  ferric  iron  rather 
stiffen  the  charge,  thus  preventing  premature  melting. 

93.     THE  ORE  HEARTH. 

The  ore  hearth  cannot,  as  regards  capacity  or  cost,  compete 
with  the  reverberatory,  nor  is  it  used  in  silver-lead  smelting. 


OF    THE    COMMON    METALS. 


309 


As  compared  with  the  reverberatory  it  can  be  quickly  started 
or  stopped  with  but  little  consumption  of  fuel,  so  that  it  serves 
well  the  purpose  of  extracting  the  lead  from  time  to  time  from 
small  amounts  of  non-argentiferous  ore  by  the  men  who  have 
mined  it  themselves. 


VERTICAL  SECTION  ON  LINE  G.  D. 


FRONT  ELEVATION. 


HORIZONTAL  SECTION  ON  LINE  A,  B. 


V s, 

FIG.  131.     AMERICAN  ORE-HEARTH. 


Fig.  131  represents  a  sectional  elevation  of  an  American  water- 
back  ore-hearth.  It  consists  of  a  crucible  e,  to  contain  the  lead, 
built  into  brickwork  n.  The  three  sides  of  the  hearth  above  the 
crucible  are  formed  by  a  cast-iron  water-cooled  jacket.  The  blast 
enters  at  a,  and  then,  through  nozzles  d,  to  the  hearth ;  g  is  the 
work-stone  and  h  a  kettle  or  pot,  placed  to  receive  the  lead,  and 


3IO  THE    METALLURGY 

kept  hot  by  a  wood  fire.  The  structure  is  surmounted  by  a  sheet 
iron  hood  or  stack  to  remove  the  fumes. 

Method  of  icorking  the  hearth. — With  the  aid  of  the  blast  a 
glowing  coal  fire  is  made,  filling  the  hearth.  Some  residue  from 
the  previous  run,  together  with  15  to  20  Ib.  of  high-grade  galena, 
not  less  than  pea-size,  is  spread  over  the  fire.  This  soon  becomes 
red  hot,  collecting  at  the  bottom  of  the  crucible.  More  ore  is 
added.  Then  the  materials  of  the  hearth  are  gently  pried  up  with 
a  bar  to  keep  the  mass  open  and  hot  throughout.  Lumps  form 
which  are  drawn  out  on  the  work-stone,  the  gray  slag  being  sep- 
arated, and  the  rich  residue  returned  to  the  hearth.  Ore  and  fuel 
are  again  added,  and  these  operations  continue  until  the  lead 
fills  the  hearth,  while  on  top  of  it  floats  the  fuel  and 
unreduced  ore.  Ore  and  fuel  are  added,  the  former  20  to  30 
Ib.  at  a  time,  sprinkled  over  the  charge.  One  man  at  intervals 
with  a  bar  loosens  and  stirs  the  charge,  raising  it  slowly,  while 
another  withdraws  upon  the  work-stone  the  semi-fused  mass  float- 
ing next  to  the  lead.  Here  he  separates  the  gray  slag,  rejecting 
it,  and  returns  the  broken-up  rich  residue  to  the  surface  of  the 
charge.  A  fresh  charge  is  then  added,  and  so  the  work  progresses. 
The  lead  as  it  accumulates  in  the  crucible  overflows,  running  down 
a  groove  made  in  the  work-stone  or  plate  to  the  kettle,  whence 
it  is  drawn  off  by  a  spout  into  molds. 

Lead  is  obtained,  as  in  the  reverberatory,  by  roasting  and  re- 
action ;  and  PbO,  when  formed,  is  reduced  at  once  to  lead  by  the 
fuel. 

The  ore-hearth  needs  power  and  a  blower,  and  much  lead  is 
volatilized,  so  that  it  is  not  suited  to  argentiferous  ores.  The 
gray  slag  which  is  obtained  still  contains  35  to  40^  Pb,  which  is 
sold  to  the  blast-furnace  smelting  works.  The  direct  recovery  of 
lead  is  no  more  than  75%. 

94.     SILVER-LEAD  SMELTING. 

This  is  a  blast-furnace  method  of  treatment,  applicable  to  the 
greatest  variety  of  ores  containing  lead,  gold,  silver  and  even  cop- 
per. By  this  method,  ores  containing  the  precious  metals,  but  no 
lead,  are  treated  with  lead-bearing  ores,  thus  using  lead  as  the 


OF    THE    COMMON    METALS. 


FIG.   132.     SILVER-LEAD  BLAST-FURNACE    (LONGITUDINAL  ELEVATION). 


312  THE    METALLURGY 

collector  of  the  values.  It  is  the  most  elective  method  known  for 
recovering  these  values,  they  being  extracted  from  the  ore  by  a 
reduction  blast-furnace  treatment  with  carbonaceous  fuel  and 
fluxes.  Oxidized  ores  (oxides  and  carbonates)  are  treated  di- 
rectly in  the  blast-furnace,  but  sulphides  are  first  roasted,  as  is 
more  fully  described  in  the  chapter  on  roasting.  This  has  become 
an  important  preliminary  operation  in  smelting. 

The  ores,  which  are  to  be  treated,  are  put  into  the  furnace  to- 
gether with  a  predetermined  quantity  of  fluxes,  taking  the  pre- 
caution of  using  enough  lead-bearing  ore  so  that  the  lead  shall 
constitute  upward  of  10%  of  the  charge.  It  has  been  found  that 
if  much  less  than  this  proportion  is  used,  the  precious-metal  con- 
tents of  the  ores  are  not  well  collected  into  the  base-bullion  or 
work-lead,  which  is  the  result  of  the  smelting.  To  the  charge 
thus  constituted  is  added  upward  of  15%  of  coke,  which  not  only 
does  the  melting,  but  also  reduces  the  lead  to  the  metallic  form 
and  the  iron  oxides  to  a  lower  state  of  oxidation. 

95.     THE   SILVER-LEAD   BLAST-FURNACE. 

Fig.  132  represents,  half  in  section,  half  in  side  elevation,  a 
view  of  a  water-jacketed  lead  blast-furnace,  rectangular  in  plan 
and  in  size  144  by  44  in.  at  the  tuyeres.  Fig.  133  is  a  half  section 
and  half  end  elevation  of  the  same  furnace.  It  consists  of  three 
parts,  the  first  being  the  shaft,  extending  from  the  feed-floor  to 
the  second  part  or  crucible,  and  surmounted  by  the  third  part,  a 
stack  or  closed  top.  Fig.  134  is  a  perspective  view  of  such  a 
furnace. 

Beginning  with  the  foundation,  we  find  it  made  of  rubble- 
masonry  or  of  concrete,  extending  from  the  solid  ground  to  the 
slag-floor  level  as  a  single  mass,  large  enough  to  sustain  the  cor- 
ner posts  also.  Where  there  is  another  furnace  close  by,  the  slag 
produced  in  this  may  with  little  expense  be  employed  for  making 
the  foundation  of  this  furnace. 

Upon  it  rests  the  crucible  shown  in  section  in  Fig.  132  and  133, 
and  in  perspective  in  Fig.  134.  In  Fig.  134  the  crucible  is  bound 
by  steel  plate.  The  bottom  of  the  crucible  consists  of  a  pan  ex- 
tending over  the  whole  area  and  including  the  exterior  binding 


OF    THE    COMMON    METALS. 


3*3 


FIG.  133.     SILVER-LEAD  BLAST-FURNACE  (TRANSVERSE  ELEVATION). 


314  THE    METALLURGY 

crucible-plates.  The  crucible,  144  by  44  in.  in  size  at  the  top,  has 
heavy  walls  and  a  bottom  of  fire-brick,  and  at  one  side  a  channel 
or  lead-well  is  built,  as  may  be  seen  in  Fig.  133.  In  operation, 
the  crucible  and  lead-well  are  full  of  lead,  and  as  fast  as  more  is 
made,  the  excess  is  removed  at  the  lead-well,  whose  opening  will 
be  noticed  in  Fig.  134.  The  jackets  form  the  lower  part  or  bosh 
of  the  shaft  and  may  be  made  of  cast  iron  or  of  steel,  as  in  Fig. 
132,  133  and  134.  The  side  jackets  have  openings  for  the 
tuyeres.  It  will  be  noticed  that  the  bosh  begins  above  the  tuyeres, 
making  a  bend  or  knee  in  the  jacket,  as  shown  in  the  cut  of  the 
steel  side-jacket,  Fig.  135,  as  well  as  in  Fig.  133.  At  the  top  of 
the  jacket  will  be  seen  a  spout  where  the  water  enters,  and,  at  the 
front  of  the  spout,  the  hole  for  the  exit  water-pipe.  Fig.  136  is 
a  steel  end- jacket  which  is  made  shorter  than  the  side- jackets  so 
as  to  leave  an  opening  below  for  the  breast. 

Jackets  are  also  made  of  cast  iron,  but  narrower  than  when  of 
steel ;  a  side- jacket,  for  example,  taking  in  but  a  single  tuyere 
and  being,  therefore,  but  1 8  to  20  in.  wide.  Each  jacket  has  its 
own  spout  for  the  inlet  and  outlet  of  water.  Hand-holes  near  the 
bottom  are  put  in  so  that  accumulated  scale  may  be  removed.  Fig. 
138  is  a  left-hand  end-jacket.  In  Fig.  132  it  will  be  seen  that 
there  is  no  bosh  in  the  end- jackets,  but  often,  as  in  Fig.  134,  such 
a  bosh  is  provided.  Fig.  138  shows  such  a  bosh,  and  indicates  the 
method  by  which  it  is  connected  to  the  side-bosh,  by  a  rounded 
corner. 

Air  is  supplied  to  the  tuyeres  from  the  bustle-pipe,  shown  in 
Fig.  133  and  134,  and  in  the  latter  figure  are  shown  the  canvas 
sleeves  by  which  the  connection  is  made.  This  makes  a  flexible 
connection  so  that  tuyeres  may  be  readily  removed,  as  when  the 
furnace  has  to  be  stopped  for  a  short  time.  A  system  of  water- 
pipes  supplies  the  jackets. 

The  main  shaft  of  the  furnace,  extending  from  the  top  of  the 
jackets  to  the  feed-door,  is  of  brick  lined  with  fire-brick,  the  whole 
being  firmly  tied  with  rods  and  angle-irons  at  the  corners.  This 
brick  structure  is  supported  by  a  deck-plate  resting  upon  the  cast- 
iron  corner-posts  or  columns  of  the  furnace. 

Above  the  feed-floor  is  the  stack  or  closed  top,  which  collects 
the  smoke  and  gases  arising  from  the  charge  and  delivers  them 


_ 

: 

FIG.  134.     PERSPECTIVE  VIEW  OF  SILVER-LEAD  BLAST-FURNACE. 


316  THE     METALLURGY 

to  the  down-take,  a  circular  pipe  of  5  ft.  diam.     This  pipe  rises 
at  an  angle  of  45°,  then  returns  at  the  same  angle  to  the  flue- 


FIG.   135.     SIDE-JACKET   WITH   KNEE-BOSH. 


FIG.  136.     STRAIGHT  END-JACKET. 

chamber.  The  stack  is  carried  up  in  brick  and  is  closed  with  a 
damper,  which  in  ordinary  operation  is  closed,  but  which  is  opened 
when  running  down  or  emptying  the  furnace. 


OF   THE    COMMON    METALS. 


317 


p  E-" 

i 

| 

i  — 

^0 

•f    t       i 

?I 

\r 

.  1 

FIG.   137.     CAST-IRON   SIDE-JACKET. 


Fie.  138.    CAST-IRON  END-JACKET. 


3l8  THE    METALLURGY 

96.     BLOWING-IN  THE   SILVER-LEAD   BLAST-FURNACE. 

This  may  be  divided  into  four  operations,  namely :  warming 
the  crucible,  melting  in  the  lead,  filling  the  furnace,  and  starting 
the  smelting. 

Warming  the  crucible. — This  is  done  gradually.  A  small  wood 
fire  with  lighter  wood  is  started  so  as  to  drive  off  the  moisture. 
This  may  take  24  hours.  The  fire  should  be  regular,  the  wood  be- 
ing placed  at  the  walls,  leaving  the  middle  of  the  crucible  clear. 
When  the  fire  has  been  kept  up  some  hours  the  charcoal  and  ash, 
both  non-conductors  of  heat,  seem  to  accumulate.  Remove  these 
and  the  glowing  coals  with  a  long  iron-handled  shovel,  and  then 
put  in  a  new  fire.  When  the  brick-work  seems  dry,  increase  the 
firing  to  warm  it  up.  To  do  this  tie  up  all  the  tuyere  sacks  and 
attach  to  the  last  tuyere  connection  of  the  bustle-pipe  a  length  of 
tuyere  sacking  long  enough  to  supply  air  to  a  6  ft.  piece  of  2l/2  in. 
pipe.  The  blower  being  slowly  rotated,  the  end  of  the  pipe,  placed 
well  down  in  the  crucible,  supplies  air  to  vigorously  consume  the 
wood.  The  lead-well,  having  a  down-draft  toward  the  crucible, 
is  warmed  at  the  same  time  \vith  \vood  and  charcoal.  Particular 
attention  should  be  given  to  heating  this  part  of  the  crucible,  and 
the  heating  is  kept  up  until  the  outer  brick-work  of  the  crucible  is 
hot  to  the  hand. 

Melting  in  the  lead. — The  crucible  is  cleaned  out  and  some  fresh 
sticks  put  in,  which  soon  take  fire.  Upon  the  wood  is  charged 
in  with  a  paddle  a  dozen  bars  of  lead  which  are  melted 
down  speedily  with  the  blast  pipe.  More  wood  is  added, 
and  other  bars,  and  thus  progressively  the  wood  is  burned 
and  the  lead  melted.  When  the  crucible  is  half  full,  the 
fuel,  charcoal  and  ashes  are  best  removed,  fresh  fire  put 
in,  and  the  melting  concluded.  It  takes  upward  of  30,000 
Ib.  of  lead  to  fill  the  crucible  of  a  42  by  120  in.  furnace.  The 
melting  begun  in  the  evening  should  be  completed  and  the  lead 
skimmed  clean  by  6  A.  M.,  when  the  day  shift  comes  on. 

Filling  the  furnace. — Dry  wood,  preferably  in  long  sticks  about 
4  in.  through,  are  put  upon  the  lead,  as  much  as  can  be  put  in  at 
the  breast.  The  tap- jacket  is  set  in  and  the  breast  bricked  up, 
while  at  the  same  time  more  wood  is  dropped  from  the  charge 


OF    THE    COMMON    METALS.  319 

door  to  fill  the  jacket  space  to  just  above  the  tuyeres.  Sound 
lump-charcoal  is  then  put  in  up  to  one-half  the  height  of  the 
jackets,  followed  by  coke  to  the  depth  of  12  to  18  in.  The  first 
charges  of  an  easy-melting  slag  follow,  with  fuel  and  flux  enough 
to  take  care  of  the  excess  silica  of  the  slag.  Ore  charge  gradu- 
ally replaces  the  slag  charge,  and  the  high  percentage  of  fuel  is 
gradually  cut  until  the  normal  charge  is  reached.  Care  must  be 
taken  that  the  furnace  is  evenly  and  quickly  filled,  extra  labor 
being  used  to  assist  in  so  doing. 

Starting  the  smelting. — The  furnace  being  half  full,  including 
some  ore  charges,  fire  is  started  at  several  of  the  tuyeres  on  either 
side  by  putting  in  a  little  greasy  waste  at  each  and  lighting  it. 
The  tuyeres  are  at  once  put  in  place  and  the  blower  started  with 
a  light  blast  of  I  or  2  oz.  pressure.  The  wood  soon  takes  fire  and 
the  smoke  rises  through  the  charge.  The  blast  is  gradually  in- 
creased during  one  or  two  hours,  and  the  furnace  as  gradually 
gets  into  operation.  As  the  slag  accumulates,  it  is  tapped  off, 
while  the  lead,  accumulating  in  the  crucible  and  lead-well,  is 
either  removed  by  ladles  or  tapped  off  by  the  lead  tap  situated 
near  the  top  of  the  lead-well.  The  amount  of  lead  removed  at  a 
time  should  be  limited  to  from  1,000  to  1,500  lb.,  and  the  lead- 
well  should  be  kept  full.  The  operation  of  filling  and  starting 
takes  about  7  hours. 

Regular  work  on  the  charge-floor. — This  consists  in  bringing 
ore,  flux  and  fuel  from  the  bins  to  the  charge  scales,  weighing 
out  the  required  amounts,  dumping  and  feeding  them  into  the 
furnace.  Everything,  except  the  foul  slag,  should  be  weighed 
at  the  furnace,  even  putting  in  the  latter  in  regular  quantity.  The 
materials  of  the  charge  weighed  in  a  charge-car  or  buggy,  are 
terials  of  the  charge,  weighed  in  a  charge-car  or  buggy,  are 
dumped  on  the  charge  plate  at  one  side  of  the  furnace,  the  coke 
for  it  being  dumped  on  the  other  side.  The  coke  is  fed  in  an  even 
layer,  and  the  plate,  thus  cleared,  receives  an  ore  charge.  The 
ore  charge  on  the  other  side  is  then  shoveled  in,  taking  care  that 
the  larger  ore  goes  to  the  middle  portion  of  the  furnace,  the  finer 
to  the  walls  and  corners.  The  blast  tends  to  creep  up  the  walls, 
more  than  at  the  middle,  but  by  this  arrangement  is  compelled 
to  come  up  evenly  both  at  the  centre  and  at  the  sides.  The  plate 


32O  THE    METALLURGY 

just  cleared  of  the  ore  charge  now  receives  coke,  which  is  the  next 
to  go  in.  The  furnace  is  maintained  at  the  level  of  the  charge- 
floor,  the  charges  being  put  in  as  the  surface  sinks. 

Regular  work  on  the  furnace  floor, — This  consists  in  regulating 
the  water  supply  to  the  jackets,  seeing  that  the  tuyeres  are  clean 
and  open,  tapping  the  slag,  and  wrhen  slag  and  matte  are 
caught  in  a  forehearth,  tapping  off  the  matte,  placing  the  slag 
pots  and  removing  them  to  the  dump ;  also  tapping  or  ladling  the 
lead  (base  bullion)  into  molds  and  removing  and  piling  the  bars 
or  ingots  to  be  sampled.  The  matte  is  removed  in  pots  or  molds, 
and  when  solidified  is  broken  up  and  sent  to  be  crushed  for  roast- 
ing. The  slag,  as  in  copper  practice,  may  be  granulated  and  swept 
away  in  a  launder  by  means  of  a  stream  of  wrater.  There  is  an 
objection  to  this,  however,  that  by  reason  of  irregular  working 
losses  of  values  may  occur  which  are  never  recovered  when  once 
the  slag  is  granulated ;  hence  removal  by  slag-carts  or  pots  is 
preferred. 

When  a  slag-pot  has  stood  for  a  few  minutes  at  the  edge  of  the 
dump,  its  contents  are  carefully  poured  out.  There  is  left  behind 
a  shell  of  solidified  slag  of  half  an  inch  or  more  in  thickness. 
This  shell  is  returned  to  the  furnace  for  re-smelting.  It  will  be 
found  to  have  within  it  drops  of  matte  which  have  settled  out  of 
the  molten  slag.  Slag  is  also  conveniently  removed  in  large  slag- 
pots,  mounted  on  trucks,  which  are  handled  by  a  horse  or  by  an 
industrial  locomotive.  At  some  large  silver-lead  smelting  works, 
the  slag  and  matte  together  are  taken  to  a  reverberatory  separat- 
ing furnace  where  the  matte  becomes  thoroughly  settled  from  the 
slag  and  is  tapped  off  at  the  bottom  level  of  the  furnace-hearths. 
The  slag  is  kept  to  the  depth  of  18  to  24  in.  in  this  furnace,  and 
is  tapped  off  near  the  surface,  when  full,  and  is  delivered  to  a 
string  of  slag-cars  handled  by  a  locomotive.  A  fire,  with  a  smoky 
reducing  flame,  is  maintained  in  this  settler.  It  gives  excellent 
results  but  needs  the  output  of  several  furnaces  to  keep  it  hot  and 
in  good  working  condition. 

97.     REACTIONS  OF  THE  BLAST-FURNACE. 

The  surface  of  the  charge  in  a  lead  furnace  should 
look  dead,  not  showing  any  overfire.  In  fact,  with  much 


OF   THE    COMMON    METALS.  321 

overfire,  there  is  loss  due  to  the  volatilizing  of  the  lead. 
The  moisture  in  the  charge  is  soon  dried  out  by  the  heated  cur- 
rents of  rising  gases.  The  heat  thus  used  is  but  small,  amounting 
to  about  one-thirtieth  of  the  total  fuel,  with  5%  moisture  in  the 
charge.  As  the  charge  descends,  CO2  begins  to  come  off  from 
the  limestone,  and  iron  is  reduced  from  the  ferric  to  the  ferrous 
form.  The  lead,  in  the  oxidized  portion  of  the  ore,  is  reduced  by 
the  CO  of  the  gases  and  especially  by  the  red-hot  coke.  It  falls 
down  in  drops  through  the  charge,  and  comes  in  contact  with  the 
ore  and  molten  slag,  taking  up  the  gold  and  silver  values  of  the 
ores.  The  sulphur  in  the  charge  takes  up  copper,  iron,  and  even 
10  to  20%  of  lead,  forming  matte.  The  iron  left  after  the  matte 
is  satisfied  goes  to  the  slag  as  FeO  and,  together  with  the  CaO, 
forms  a  fusible  silicate.  The  molten  products  separate  at  the 
hearth  according  to  their  specific  gravity,  that  of  lead  being  11.5, 
that  of  matte  averaging  5.2,  and  that  of  slag  3.6.  The  lead  col- 
lects in  the  crucible  and  is  withdrawn  at  the  lead-well.  The  other 
products,  matte  and  slag,  are  drawn  off  at  the  slag-tap,  situated 
just  above  the  level  of  the  lead.  The  separation  of  the  matte  from 
the  slag  is  generally  effected  in  a  forehearth  outside  the  furnace. 

98.     LEAD  SLAGS. 

The  object  of  silver-lead  smelting  is  to  separate  the 
contained  lead  from  the  ore;  the  sulphur  present  forming,  with 
copper  and  iron,  some  matte ;  while  the  remaining  available 
bases,  united  to  the  silica,  form  a  slag.  If  there  are  not  bases 
enough  to  form  a  suitable  slag,  they  must  be  added  in  the  form 
of  fluxes  to  the  charge. 

Various  authors,  in  describing  the  composition  of  slags,  have 
divided  them,  according  to  the  chemical  composition,  into : 

Singulo-silicates  (2  RO  +  SiO2),  in  which  the  ratio  of  oxygen 
in  the  bases  to  oxygen  in  the  silica  is  as  I  to  I  ; 

Sesqui-silicates  (4  RO  +  3  SiCX),  with  an  oxygen  ratio  of 
2  to  3; 

Bi-silicates  (RO  +  SiO2),  having  an  oxygen  ratio  of  I  to  2. 

Let  us  make  the  RO  bases  to  be  in  the  proportion  of  their 
equivalents,  namely:  72  parts  Fe(Mn)O  to  56  parts  Ca(Mg,Ba) 


THE    METALLURGY 


O,  then,  with  the  total  of  three  elements  forming  90%  of  the  slag 
(the  other  bases  are  assumed  to  be  10%)  we  have: 

Singulo-silicates  28.7%  SiO2,  34.5%  Fe(Mn)O,  and  26.8% 
Ca(Mg,  Ba)O=90%. 

Sesqui-silicates,  $7.2%  SiCX,  29.8%  Fe(Mn)O,  and  23.0% 
Ca(Mg,  Ba)O 


Bi-silicates,  43.6%  SiO2,  26.1%  Fe(Mn)O,  and  20.3%  Ca 
(Mg,  Ba)O=90%. 

However,  in  practice,  this  way  of  looking  at  the  constitution  of 
slags  is  seldom  followed  since  slags  are  silicates  of  considerable 
complexity,  but  the  percentages  of  them  are  what  is  generally 
given. 

Typical  slags,  as  those  which  have  proportions  of  SiO2, 
Fe(Mn)O  and  Ca(Ba,  Mg)O,  such  as  are  suited  for  successful 
use  in  the  blast-furnace.  To  fulfill  the  requirements  of  a  good 
slag,  they  should,  in  the  normal  operation  of  the  furnace,  con- 

Table  of  Typical  Slags. 


Type 

SiO2 

* 

Fe  (Mn)O 
jf 

Ea(Ba,Mg)O 
* 

Quarter-slag 

C             28 

50 

12 

Silicious  quarter-slag 

H            32 

47 

11 

Half-slag 

E            30 

40 

20 

Half-slag 

J             31 

38 

21 

Silicious  half-slag 

I              35 

38 

17 

Three-quarter-slag 

F             33 

33 

23 

Silicious  three-quarter-slag 

M             36 

31 

23 

Whole,  or  1  to  1  slag 

G             35 

27 

28 

I 

tain  not  more  than  0.7%  Pb,  nor  should  the  Ag  be  more  than 
0.5  oz.  when  the  charge  is  producing  base  bullion  not  higher  than 
300  oz.  per  ton ;  neither  should  they  have  a  density  greater  than 
3.6,  nor  permit  accretions  at  the  hearth  or  the  creeping  up  of 
overfire. 


OF   THE    COMMON    METALS.  323 

According  to  the  ratio  of  FeO  to  CaO,  so  slags  are  called  a 
quarter,  a  half,  a  one  to  one,  etc.,  slag.  Thus  slag  E  of  the  sub- 
joined table  is  called  a  half-slag,  the  CaO  being  but  one-half  the 
percentage  of  the  FeO.  Slag  C  is  a  quarter  slag,  the  CaO  being 
approximately  one-fourth  of  the  FeO.  In  this  table  of  typical 
well-tested  slags  the  three  elements  are  reckoned  at  90%.  If  the 
sum  should  vary  from  this,  the  ratio  may  be  still  preserved. 

Action  of  elements  of  the  slag. 

Iron. — Iron  ore  is  soon  reduced  to  the  ferrous  form  under  the 
action  of  carbon,  thus,  Fe2O3  +  C  =  2  FeO  +  CO,  when,  being 
the  stronger  base,  it  replaces  lead  oxide  in  silicious  combination, 
the  lead  oxide  then  being  reduced  to  metallic  form,  and  finding  its 
way  as  base  bullion  or  work-lead  to  the  crucible,  while  the  iron 
serves  as  a  base  for  the  silica,  thus : 

Pb  SiO3  +  FeO  +  C  =  Fe  SiO3  +  Pb  +  CO 

Galena  is  also  reduced  to  metallic  lead  in  presence  of  ferrous 
iron  and  carbon  as  follows  : 

PbS+FeO+C=rFeS+Pb+CO.  This  reduction  is  by  no 
means  perfect.  A  little  PbS  with  whatever  copper  may  be  in  the 
charge,  and  the  needed  iron  and  sulphur,  forms  that  complex 
copper-lead-iron  sulphide  called  'matte/ 

Manganese.  The  equivalent  for  manganese  is  55,  that  of  iron 
56,  so  that  we  may  reckon  them  as  having  equal  values  for 
fluxing.  . 

Lime,  magnesia  and  baryta.  These  alkaline  earths  act  in  in- 
verse ratio  of  their  atomic  weights  in  fluxing  silica.  Hence,  to 
obtain  the  equivalent  lime,  their  percentages  are  multiplied  by  1.4 
for  magnesia  and  by  0.4  for  baryta.  Lime,  replacing  FeO,  tends  to 
make  a  slag  of  less  specific  gravity,  this  property  causing 
a  better  separation  of  slag  from  matte.  Being  a  stronger  base 
than  iron,  and  generally  a  cheaper  flux,  the  tendency  is  to  substi- 
tute it  for  iron.  It  will  be  noticed  that  the  higher  the  slags  of  the 
table  are  in  silica,  the  higher  they  are  in  lime,  or  that  a  high  silica 
needs  a  high  lime.  Accordingly,  where  the  making  of  silicious 
slags  is  the  most  profitable,  there  will  the  slags  be  limy.  Dolomite 
is  avoided  for  use  in  silver-lead  smelting,  since  it  makes  the  slag 
pasty  and  streaky,  and  its  unfavorable  effect  is  shown,  especially 


324  THE    METALLURGY 

in  presence  of  zinc.  The  analyses  of  limestone  and  of  dolomite, 
given  below,  show  the  composition  of  those  fluxes  in  actual  prac- 
tice. 

Canyon  City,  Colo.,  Limestone. — 49.8%  CaO,  3%  MgO,  3.1% 
SiO2,  0.8  %  Fe. 

Iron  County,  Mo.,  Dolomite.— 26.6%  CaO,  17.6%  MgO,  5.1% 
Si02,  3.3%  Fe. 

Fluorspar  has  no  unfavorable  effect  on  the  slag;  still,  when 
present,  the  fluorine  uses  up  CaO,  and  hence  the  slag  must  have 
more  CaO  than  the  type-slag  calls  for. 

Alumina.  It  is  a  question  whether  alumina  acts  as  an  acid  or 
as  a  base.  It  will  be  sufficient  for  the  purposes  of  silver-lead 
smelting  to  regard  it  as  a  constituent,  dissolving  in  the  slag,  and 
acting  in  neither  way. 

Copper.  Where  copper  exists  in  the  ores,  it  enters  the  matte 
and  generally  such  matte  is  to  be  provided  for  it  to  enter.  How- 
ever, in  smelting  carbonate  or  oxidized  ores  without  the  forma: 
tion  of  matte,  the  copper  becomes  reduced,  and  enters  the  base 
bullion  giving  a  drossy  lead,  sometimes  so  much  so  that  it  tends 
to  clog  and  stop  up  the  lead-well  and  to  accumulate  and  solidify 
in  the  crucible.  The  remedy  is  to  add  sulphides  in  quantity  suffi- 
cient to  form  some  matte. 

Zinc.  Blende  and  zinc  oxide  cause  difficulties  in  the  blast-fur- 
nace. Blende  is  in  part  decomposed  into  zinc  oxide,  but  in  either 
form  stiffens  the  slag  and  tends  to  make  it  infusible.  It  may  be 
regarded  as  being,  like  alumina,  dissolved  in  the  slag.  It  goes  both 
into  the  slag  and  into  the  matte,  thus  lessening  their  difference  in 
specific  gravity  and  causing  a  less  perfect  separation  of  the  two. 
Where  there  is  much  zinc  in  the  charge,  it  has  been  customary  to 
modify  the  type-slag  by  figuring  one-half  of  the  zinc  as  replacing 
lime.  Thus,  in  the  half  slag  /  of  the  table,  and  where  there  is  8% 
zinc  oxide,  we  should  have  for  a  correct  slag  31%  SiO2,  38% 
FeO,  17%  CaO,  and  8%  ZnO,  making  in  all  94%.  Reducing 
these  percentages  proportionately  to  a  total  of  90%,  we  have  ap- 
proximately 29.5%  SiO2,  36.0%  FeO,  16.0%  CaO  and  7.5%  ZnO. 

Antimony.  As  a  sulphide,  this  is  reduced  like  lead,  going  into 
the  base  bullion  and  making  it  hard.  It  has  to  be  removed  later 
in  refining. 


OF   THE    COMMON    METALS.  325 

Arsenic  occurs  frequently  in  silver-lead  smelting.  It  takes  up 
iron  and  flows  from  the  furnace  as  a  'speiss.'  Where  present,  iron 
must  be  calculated  for  it.  In  the  fire  assay  of  lead  ores  for  lead 
it  will  be  found  attached  to  the  lead  button.  If,  from  the  data 
thus  obtained,  we  know  the  weight  of  speiss,  we  may  figure  that 
70%  of  it  is  iron.  When  a  direct  determination  of  the  arsenic  is 
made,  its  weight,  multiplied  by  2.3,  will  express  the  quantity  of 
Fe  which  must  be  provided  for  in  the  charge  calculation. 

Fuels  used  in  the  blast-furnace.  The  fuels  used  in  silver-lead 
smelting  are  coke,  charcoal,  or  a  mixture  of  the  two. 

Coke.  Coke  of  all  kinds  is  used,  the  ash  varying  from  10  up  to 
22%,  and  the  fixed  carbon  from  89  to  77%.  In  the  high-ash  coke 
not  only  is  more  waste  material  to  be  smelted,  but  the  carbon  is 
lower,  so  that  such  cokes  are  less  efficient.  Perhaps,  however,  the 
greatest  trouble  with  these  high-ash  cokes  is  that  they  are  often 
friable,  making  fines  and  dust,  and  obstructing  the  passage  of  the 
blast. 

Analyses  of  two  such  cokes  give : 

Connellsville  "coke,  87.5%  fixed  carbon,  11.3%  ash,  0.7%  S. 

El  Moro  coke,  77.0%  fixed  carbon,  22.0%  ash,  0.9%  S,  when 
made  from  unwashed  coal. 

In  computing  a  charge  the  coke-ash  is  taken  into  account, 
analyses  being  as  follows : 

Ash  of  Connellsville  coke,  44.6%  SiO2,  15.9%  Fe,  7.0%  CaO, 
1.9%  MgO. 

Ash  of  El  Moro  coke,  84.5%  SiO,,  5.0%  Fe. 

Charcoal.  This  fuel  is  used  where  the  cost  of  coke  would  be 
too  high.  It  is  a  good  fuel  for  oxidized  ores,  but  is  friable,  mak- 
ing undesirable  fines.  It  makes  a  more  open  charge  than  coke, 
and  contains  but  little  ash — less  than  2%.  Coke  may  be  taken 
at  40  Ib.  and  charcoal  at  13  Ib.  to  the  cu.  ft. ;  the  weight  of  char- 
coal is  from  14  to  16  Ib.  to  the  bushel.  Where  charcoal  is  the 
cheaper  fuel,  it  often  helps  the  operation  of  the  furnace  to  use  a 
proportion  of  coke,  which,  fed  at  the  walls,  and  burning  more 
slowly  than  the  charcoal,  makes  the  zone  at  the  tuyeres  hotter, 
and  gives  better  fused  slags. 

Quantity  of  coke.  This  varies  according  to  the  nature  of  the 
charge  from  12  to  16%,  or  even  more.  Charges  which  make  con- 


326  THE    METALLURGY 

siderable  matte  need  less  than  oxidized  ores.  So  much  must  be 
used  as  will  give  good  reduction  and  a  hot  slag,  and  the  metallur- 
gist is  guided  by  these  factors  in  adding  coke. 

99.     CALCULATION  OF  A  SILVER-LEAD  BLAST-FURNACE  CHARGE. 

First  example:  To  determine  the  weight  of  fluxes  to  add  to 
the  charge,  it  is  necessary  to  have  the  analyses  of  the  ores,  fluxes 
and  fuel  to  be  used,  and  to  specify  the  composition  of  the  desired 
slag  and  matte.  Below  is  given  an  example  of  a  charge-sheet  for 
a  charge,  containing  no  roasted  ore,  and  to  weigh  approximately 
1,000  Ib.  Lead  ore  must  be  added  in  quantity  sufficient  to  make 
the  lead  equal  to  10%  of  the  charge  or  100  Ib.  To  this  may  be 
added  silicious  ore,  smelted  to  obtain  its  contained  gold  and  silver. 
The  quantities  of  the  fluxes,  as  experience  suggests,  is  now  writ- 
ten in.  Let  us  estimate,  that  of  the  sulphur,  20%  is  volatilized, 
80%  goes  into  the  matte  and  some  enters  the  slag  to  the 
extent  of  i%  of  its  weight.  The  weight  of  slag  will  be 
203-^-0.30  or  677  Ib.  Neglecting  fractions,  we  have  sulphur 
entering  the  slag  7  Ib.,  sulphur  volatilized  (20^  of  23  Ib. ) 
5  Ib.,  or  a  total  of  12  Ib.,  leaving  n  Ib.  to  enter  the  matte. 
Multiplying  this  quantity  by  2.75,  gives  us  30  Ib.  of  Fe  needed  for 
the  matte.  This  leaves  201 — 30=171  Ib.  to  enter  the  slag. 
Multiplying  the  SiO2 '203  Ib.  by  1.04  gives  211  Ib.  of  Fe  needed. 
We  have,  however,  only  171  Ib.  and  the  difference  must  be 
made  up  by  increasing  the  Fe  by  40  Ib.  or  the  iron  ore  by  80  Ib. 
since  the  iron  in  the  iron  ore  is  approximately  one-half  the  ore. 
Multiplying  the  silica,  203  Ib.,  by  0.67  gives  us  131  Ib.  of  CaO 
needed,  or  we  have  36  Ib.  CaO  too  much.  Since  the  limestone  is 
approx  mately  half  CaO,  this  will  diminish  the  limestone  by  about 
70  Ib.  Erasing  the  old  values  and  substituting  the  new  ones  \ve 
carry  turough  all  the  calculations  again.  The  results  should  be 
correct  to  within  a  few  pounds,  and  are  again  corrected  if  needed. 
Fractions  of  pounds  are  to  be  neglected,  and  the  weights  of  fluxes 
need  not  be  written  in  nearer  than  the  nearest  10  Ib.,  since  varia- 
tions in  the  charge,  due  to  variations  of  the  ores  and  of  the  weigh- 
ing, would  be  greater  than  this.  It  must  be  remembered  that  the 
slag  actually  produced  will  generally  vary  from  the  calculated 


OF   THE    COMMON    METALS. 


327 


proportions.  When  a  charge  has  been  put  on,  and  after  several 
hours  the  slag  comes  down  or  begins  to  flow  from  the  furnace,  it 
should  be  quickly  analyzed  and  the  charge  corrected  by  a  suitable 
change  in  the  fluxes.  The  total  weight  of  the  charge  can  be  varied 
by  varying  the  silicious  ore  while  at  the  same  time  retaining  the 
lead  unchanged. 

Charge  Sheet. 


Weight 

Pb 

SiO, 

Fe&Mrt     Ca° 

S 

Ag 

&MgO 

H2O 

Wet 

Dry     * 

Wtj    4 

Wt 

* 

Wt 

* 

Wt. 

* 

Wt. 

Oz. 

Chrysolite  No.  28 

3.0 

500 

21.0 

105 

32.0 

160 

15.  C 

75 

10.0 

50 

4.4 

22.0 

75 

18.7 

Ontario 

5.0 

50 

60.0 

30 

2.C 

1 

2.0 

1 

0.5 

75 

1.9 

Iron  Ore 

2.0 

200 

4.0 

1 

57.0 

114 

3.0 

6 

Limestone 

200 

3.0 

6 

3.5 

7 

54.0 

108 

Coke 

150 

4.0 

6 

2.6 

4 

1.5 

2 

0.7 

1.1 

950 

105 

203 

201 

167 

23.1 

20.6 

30 

For 
Matte 

12-0  For  Slag  and 
....  _  Volatilized 

171  For  Slag 

n-°  For  Matte 

Slag  SiO2  =  30.0  =1.00      factor 

FeO  =  40.0  =  31.1  Fe  =  1.04 
CaO  =20.0  =0.67 

90.0 


Matte  S  =  20.0 
Fe  =  55.0 
Cu=  0.0 


FeO 


=  2.75  factor 


Second  example:  Where  sulphur-bearing  oxidized  and  silicious 
ores  are  to  be  used,  we  have  to  consider  not  only  these  constitu- 
ents, but  also  the  out-products  of  the  furnace. 

Ores  containing  less  than  10  to  12%  sulphur  are  generally 
smelted  without  roasting.  It  is  cheaper  to  do  this  since  in  roast- 
ing the  sulphur  would  be  removed  only  down  to  3  or  4%  anyway. 
Such  unroasted  ore  would  make  40%  of  its  weight  in  matte,  but 
that  matte,  on  roasting,  would  again  yield  20%  of  matte  or  8% 
figured  on  the  original  ore,  or  equivalent  to  the  amount  which 
would  be  produced  by  a  thoroughly  roasted  ore.  Again,  many 
ores  below  the  above  specified  limit  (pure  galena  has  13.4%  S), 
are  leady  ores  and  accordingly  quite  difficult  to  roast  because  of 
their  fusible  nature.  Ores  intended  for  roasting  may  be  simple, 
consisting  of  iron  sulphide ;  and  complex,  as  shown  by  the  follow- 
ing analysis  of  a  roasted  ore :  10.0%  SiO2 ;  27.0%  Fe  and  Mn  ; 


328  THE    METALLURGY 

2.0%  CaO,  MgO  and  BaO;  8.8%  Zn;  0.4%  Cu,  6.0%  S;  35.o7c 
Pb,  and  having  50  oz.  Ag  per  ton,  the  various  bases  having  been 
present  as  sulphides. 

The  so-called  'oxidized'  ores  are  those  consisting  of  oxides  and 
carbonates  of  lead,  with  a  gangue  of  iron  oxide,  limestone,  dolo- 
mite and  silica.  The  lime,  magnesia  and  iron  are  useful  for  flux- 
ing the  silica  of  the  charge.  Such  ores,  though  called  oxidized, 
often  carry  a  little  sulphur,  both  as  sulphides  (galena  and  pyrite), 
and  also  as  sulphates. 

Silicious  ores  are  added  to  a  charge,  in  spite  of  the  large  excess 
of  silica  which  they  carry,  because  they  have  values  in  gold  and 
silver  which  must  be  obtained  from  them.  The  lead  of  the  charge 
will  take  up  this  contained  gold  and  silver  from  such  ores,  while 
their  silicious  gangue,  fluxed  into  a  barren  slag,  is  sent  to  waste. 

Both  iron  ore  and  limestone  are  added  to  a  charge  for  fluxing 
the  silica  with  the  formation  of  a  slag  of  a  predetermined  com- 
position or  type.  If  these  fluxes  carry  any  silver  and  gold  such 
values  can  be  recovered,  they  going  into  the  base  bullion  or  work- 
lead.  Otherwise,  they  are  what  is  called  'barren'  or  'dead'  flux. 
Ores  carrying  an  excess  of  iron  or  of  lime  over  silica  (called  iron 
or  lime  excess)  are  in  the  same  category,  since  that  excess  is 
useful  as  flux  and  is  allowed  and  paid  for  in  the  purchase  of  the 
ores. 

Not  all  the  slag  which  issues  from  the  furnace  is  quite  clean. 
At  the  spout,  where  the  matte  runs  over  it,  and  in  the  shell  form- 
ing the  cavity  of  a  forehearth,  slag  containing  drops  of  lead  is  to 
be  found.  When  a  slag-pot  is  poured  out  at  the  edge  of  the  dump, 
there  remains  a  shell  or  coating  of  the  solidified  slag.  This  shell, 
half  an  inch  thick,  will  be  found  to  contain  drops  of  matte  or  of 
lead  which  has  not  entirely  settled  out  in  the  forehearth.  All  this 
value-containing  slag,  called  'foul'  slag  is  an  acceptable  addition 
to  the  charge,  because  of  its  easy  fusibility  and  its  coarse  condi- 
tion permitting  freer  passage  of  the  air  of  the  blast. 

Finally  we  have  the  fuel  constituting  say  15%  of  the  total 
charge,  the  word  'charge'  meaning  the  ores,  fluxes  and  the  foul 
slag  going  once  more  through  the  furnace. 

In  computing  a  charge,  we  use  the  subjoined  charge  sheet  as 
follows :  The  size  of  the  charge  to  be  approximately  1,000  Ib.  Let 


OF    THE    COMMON    METALS. 


329 


us  say  that  we  are  supplied  with  ore  in  the  ratio  of  the  quantities 
specified  in  the  first  three  items  of  the  charge  sheet:  roasted  ore 
300  lb.,  lead  ore  200  lb.,  and  silicious  ore  200  Ib.  Write  in  with 
a  lead  pencil  the  quantity  of  fluxes  necessary,  as  near  as  may  be 
judged  by  experience.  First  find  if  there  is  about  10%  of  lead  on 
the  charge,  the  quantity  preferably  used  for  ensuring  a  good  col- 
lection of  the  values  in  the  ore.  The  ores  containing  it  must  be 
varied  to  accomplish  this.  We  here  find  104  lb.  lead  which  meets 
the  requirements.  Next  write  in  150  lb.  coke,  or  15%  of  the 
charge,  estimating  its  percentage  composition  from  the  following 
data,  for  example,  of  Connellsville  coke:  with  11%  ash  containing 
44.6%  SiO2,  22.7%  Fe,O3,  7%  CaO,  1.9%  MgO.  Estimat- 
ing this  for  the  coke,  we  have  4.9%  SiO2,  1.7%  Fe,  1.1%  CaO 
and  MgO.  Since  the  iron  is  in  ferric  form,  we  have  to  ob- 
tain 0.7  of  the  quantity  given  to  reduce  it  to  the  equivalent 
iron.  Since  magnesia  is  a  stronger  base  than  lime,  we  multiply  its 
quantity  by  1.4  or  in  the  ratio  of  MgO  to  CaO,  adding  it  then  to 
the  CaO.  Write  in  all  the  other  percentages  and  carry  them  out 

Charge  Sheet. 


1- 

Name  of  Ore 

Weight 

Pb. 

SiO2 

Fe  &  Mn 

CaO 
&MgO 

S 

H2O 

Wet 

Dry 
300 

% 
12.0 

Wt. 

36 

« 

12.0 

Wt. 

* 
32.0 

wt 

96 

$ 
2.0 

Wt. 
6 

* 

5.0 

Wt. 
15 





Roasted  Ore  . 

48 

Lead  Ore 

200 

25.8 

52 

22.6 

45 

18.1 

36 

5.4 

11 

2.0 

4 

Silicious  Ore 

200 

8.0 

16 

63.3 

127 

16.0 

32 

3.0 

6 

Iron  Ore 

100 

5.0 

10 

54.0 

54 

Limestone 

200 

3.5 

3 

52.0 

104 

Coke 

250 

1000 

4.9 
104 

7 

1.7 

3 

1.1 

2 
123 

0.7 

1 
26 

240 

221 

Fe  for  Matte  =    37 
Fe  for  Slag     =  183 

S  volati- 
lized        15.0 
S  in  Slag  16.0 

11 

-ic    For 
!Matte 

Slag 

SiO,  =  33* 

FeO  =  33$     Fe  =  25.7  =  0.77  x  SiO, 

CaO  =  24$  =  0.73  x  SiO, 

Other  Bases  =  10$ 
100$ 

Slag  -  730  lb. 
S  in  Slag  =  (730  x  0.8$)  =  6  lb. 


Matte 

S     =20* 

20$  S  x  2.5  =  Fe  =  50$ 
Cu=  5* 
Pb  -  15* 


33O  THE    METALLURGY 

in  pounds,  footing  up  the  columns  to  obtain  the  totals  of  each 
constituent.  All  fractions  are  to  be  neglected.  Put  down  at  the 
left  side  of  the  charge  sheet  a  type  of  slag  such  as  will  be  found 
in  the  list  of  type-slags,  section  99.  The  type  to  be  chosen  will 
depend  upon  what  ores  it  pays  best  to  treat,  and  upon  the  cost 
of  fluxes.  When  it  pays  best  to  treat  silicious  ores,  one  would 
use  a  silicious  slag ;  but  when  there  is  more  profit  in  an  irony  ore, 
then  use  a  basic  slag.  For  this  calculation  we  will 
use  a  three-quarter  slag  of  33%  SiO2,  33^  FeO  and 
MnO,  24%  CaO  and  MgO,  while  the  other  bases,  A12O3, 
ZnO,  alkalies,  and  S  will  make  up  the  remaining  10% .  The  FeO 
(33%)  ls  calculated  to  Fe,  making  it  25.7^.  Find  the  ratio  of 
SiO2  to  Fe,  which  will  give  us  the  factor  0.77.  For  CaO  we  will 
in  the  same  way  get  the  factor  0.73.  On  the  other  side  of  the 
sheet  put  down  the  matte  analysis  and  calculate  the  factor  express- 
ing the  ratio  of  S  to  Fe,  or  2.5. 

Let  us  first  consider  the  sulphur.  Experience  shows  that  25% 
or  one-fourth  of  it  is  volatilized,  also  that  the  slag  contains  0.8% 
S.  The  slag  itself  is  obtained  by  dividing  the  240  Ib.  of  SiO0  by 
its  percentage  (33)  and  equals  730  Ib.  Hence  there  remains  15 
Ib.  sulphur  to  enter  the  matte,  which  needs  15X2.5  or  37  Ib.  Fe. 
This,  subtracted  from  the  total  Fe,  will  leave  183  Ib.  for  the  slag. 
But,  by  calculation,  for  the  240  Ib.  of  SiO2  we  shall  need  (24oX 
0.77)  or  185  Ib.  Fe,  which  is  near  enough.  Were  it  desirable  to 
correct  this  small  amount,  we  proceed  thus.  The  Fe  being  ap- 
proximately one-half  of  the  iron  ore,  then  for  2  Ib.  lacking  of  Fe 
we  shall  need  4  Ib.  iron  ore  and  this  would  be  the  new  corrected 
quantity.  Again,  calculating  the  lime  needed  (240  by  0.73)  or 
175  Ib.,  we  find,  by  comparing  with  the  actual  amount,  that  we 
are  short  (175 — 123)  52  Ib.,  equivalent  to,  say,  no  Ib.  of  lime- 
stone of  about  50%  CaO.  Xeatly  amend  the  figure  needing 
change  in  the  limestone  column,  using  the  new  figure  of  310  Ib. 
for  that  item  of  the  charge.  This  will  alter  the  footings  of  the 
columns,  and  the  older  calculations,  all  of  which  must  be  neatly 
erased  and  changed  as  required.  The  amended  calculation  will 
be  nearly  correct,  but  if  not  so,  make  a  new  correction  with  the 
older  figures  erased,  and  the  new  ones  put  in.  The  final  sheet 
should  be  a  fair  copy. 


OF   THE    COMMON    METALS.  331 

In  conclusion  we  will  note  that  since  the  equivalents  of  Fe  and 
Mn  are  so  near  one  another,  we  may  add  their  respective  per- 
centages for  the  equivalent  Fe.  For  magnesia  multiply  by  1 .4  and 
for  baryta  by  0.4  to  obtain  the  equivalent  CaO. 

100.     HANDLING  BASE  BULLION. 

The  practice  in  large  works  is  to  re'melt  all  the  work- 
lead  or  base  bullion  from  the  blast-furnace.  The  skim- 
mings from  this  operation  are  returned  to  the  furnace,  while 
the  cleaned  base  bullion  is  molded  into  bars.  While  molding, 
samples  are  taken  from  the  melting  kettle,  from  which  the  assay 
results  are  obtained.  The  bars  (400  to  500  in  the  carload)  are 
stamped  with  the  number  of  the  lot  and  are  carefully  weighed, 
20  at  a  time.  A  careful  assay  is  made  upon  each  lot,  both  by  the 
shipper  and  by  the  refinery.  In  smaller  works  the  punch  sample, 
already  described,  is  taken. 

101.     FLUE-DUST  FROM  THE  SILVER-LEAD  BLAST-FURNACE. 

A  standard  size  blast-furnace,  42  by  120  in.  may  take  5,000  cu. 
ft.  of  air  when  in  full  operation.  This  is  expanded  by  the  heat 
of  the  furnace  at  least  twice  this  volume,  and  by  taking  up  the 
fuel  in  gaseous  form,  so  that,  in  escaping  through  the  interstices 
of  the  charge,  strong  upward  currents  result.  Thus,  fine  parti- 
cles of  ore  are  carried  by  the  down-take  to  the  flues.  These  flues, 
of  large  cross  section,  may  be  hundreds  of  feet  in  length  for  the 
purpose  of  thoroughly  settling  out  and  collecting  these  particles, 
which  are  subsequently  made  into  briquettes  and  re-smelted.  Be- 
sides this,  a  portion  of  the  lead,  silver,  and  zinc  contained  in  the 
ore,  may  be  volatilized,  and  this,  when  cooled  in  the  flue,  may 
deposit,  at  least  in  part.  Thus,  flue-dust  may  be  considered  as 
composed  of 

Dust,  carried  along  by  the  draft. 

Lead-fume,  condensed  on  cooled  surfaces. 

All  this  material  is  not  saved,  but  the  finest  part  may  escape 
together  with  sulphur  dioxide  and  arsenic.  The  latter,  however, 


33^  THE    METALLURGY 

condenses  much  like  the  lead,  and  where,  as  in  some  out-lying 
place,  the  lead  has  a  lower  value,  there  may  be  a  point  where,  in 
order  to  get  rid  of  arsenic,  it  may  be  profitable  to  permit  such 
loss.  There  is  a  great  difference  in  the  dust-loss  according  to  the 
way  the  furnace  runs.  A  furnace  having  much  shaft  accretions, 
one  driven  with  much  volume  of  blast,  or,  where  the  feeding  is 
carelessly  performed,  all  gives  an  increased  flue-dust  loss.  To 
catch  as  much  as  possible  of  the  dust,  the  method  which  has 
proved  most  satisfactory  is  to  have  flues  of  large  sectional  area, 
in  which  the  velocity  of  the  gases  is  made  low.  In  this  quiet  place 
the  dust  settles  to  the  best  advantage.  Flues  have  also  been  made 
of  sheet  metal,  but  they  corrode  under  the  influence  of  the  sul- 
phuric acid  and  sulphates  formed  in  the  furnace  and  condensed  on 
the  metal.  Reinforced  concrete  has  been  used,  but  is  also  some- 
what attacked,  so  that  brick  remains  the  favorite  material  for  such 
construction.  A  flue  built  on  the  Monier'  system  consists  of  a 
framework  of  arched  angle-iron  tied  by  longitudinal  bars  and  cov- 
ered by  wire  netting  or  expanded  metal  and  the  whole  plastered 
over  inside  and  out  with  a  cement  concrete.  The  bottom  of  brick 
flues  is,  however,  frequently  made  of  metal  in  hopper  form.  Since 
these  surfaces  are  covered  by  the  flue-dust  they  are  protected  from 
corrosion  and  last  well.  A  cross-section  of  such  a  flue  shows 
a  hopper-bottom  under  which  runs  a  track.  A  covered 
tramcar  can  then  be  set  under  any  hopper  and  its  contents  drawn 
into  the  car  through  a  canvas  sleeve  fitted  upon  the  outlet  spout, 
thus  avoiding  dusting.  Other  methods,  such  as  water  sprayed 
upon  the  dust-laden  air,  plates  or  wires  hung  in  the  flue  (Freu- 
denberg  plates,  Rosing  wires)  have  been  tried  with  some  degree 
of  success,  but  have  been  given  up  in  favor  of  plain  flues  without 
bafBle-walls  or  other  obstructions  in  the  flue.  The  two  points  to 
consider  are  first,  a  slow  and  gentle  movement  of  the  gases  to 
settle  out  the  dust,  and,  second,  large  cooling  surfaces  for  the  con- 
densation of  the  lead  fumes. 

102.     PREPARATION  OF  FLUE-DUST  FOR  RE-SMELTING. 

Flue-dust  can  be  wet  down  and  fed  back  to  the  blast-furnace. 
If  returned,  a  little  at  a  time,  it  is  simply  blown  back  into  the  flue, 


OF    THE    COMMON    METALS. 


333 


but  it  may  be  fed  while  wet  in  occasional  large  charges,  so  that 
most  of  it  is  smelted.  The  proper  way,  however,  is  to  make  it  into 
briquettes,  with  milk-of-lime  for  binder.  Fig.  139  represents  a  plant 
containing  a  White  briquetting  press  for  making  such  briquettes 
(composed  of  flue-dust  and  milk-of-lime  to  which  is  often  added 


FIG.  139.     WHITE  BRIQUETTING  PRESS. 


fine  roasted  ore).  At  the  right,  on  the  higher  platform,  is  to  be 
seen  a  pile  of  burned  lime.  This  is  fed,  together  with  water,  into 
the  mixer,  and  there  slacked  into  a  thin  paste,  which  is  allowed 
to  constantly  run  into  the  near  end  of  the  pug-mill  O.  Flue-dust, 
from  the  pile  at  the  front  of  the  lower  platform,  is  shoveled  into  O 


334  THE    METALLURGY 

and  becomes  thoroughly  mixed  by  the  revolving  knives  of  the 
pug-mill  which,  set  at  an  angle,  screw  it  along  to  the  discharge 
opening  immediately  over  the  troughed  conveying  belt.  It  drops 
into  the  hopper  of  a  six-mold  briquetting-press,  where  it  is  made 
into  briquettes  under  strong  pressure,  and  then  comes  upon  the 
belt  i,  which  delivers  it  into  a  heap.  The  briquettes  may  be  used 
up  at  the  blast-furnace  as  made,  or  may  be  allowed  to  dry  and 
harden  before  use. 


103.     TREATMENT  OF  LEAD-COPPER  MATTE 

The  shipping  matte  from  the  treatment  of  lead-silver  ores  runs 
as  high  as  45  to  5o9r  in  copper.  This  matte  may  be  crushed  to 
4-mesh,  and  roasted  in  the  ordinary  long-bedded  reverberatory 
furnace.  It  is  then  smelted  in  a  blast-furnace  together  with  some 
silicious  and  oxidized  copper  ore.  There  results  a  matte  of  65% 
Cu,  together  with  a  considerable  proportion  of  bottoms,  the  result 
of  the  reduction  of  part  of  the  copper  from  the  matte.  These  bot- 
toms contain  the  gold,  lead  and  the  impurities,  such  as  As,  Bi 
and  Sb. 

The  bottoms  are  now  charged  into  a  reverberatory  furnace 
through  the  side  doors.  The  coarsely  broken  matte  is  then  put  in 
on  top,  the  doors  are  closed,  and  the  charge  is  slowly  roasted,  as 
in  the  Welsh  process  of  producing  blister  copper.  The  charge 
having  been  melted  down,  the  reaction  of  the  cuprous  oxide  on 
the  cuprous  sulphide  sets  in  until  the  charge  is  reduced  to  copper. 
It  is  then  poled,  to  free  it  from  copper  oxide,  and  is  ladled  into 
anodes. 

104.     LEAD  ORES — COST  OF  TREATMENT. 

To  illustrate  the  method  of  calculating  the  actual  cost  of  treat- 
ing an  ore  in  Western  silver-lead  practice,  we  may  take  a  so-called 
neutral  ore  (Fe  equals  SiO2)  as  follows:  It  is  assumed  to  con- 
tain less  than  5%  sulphur  and  to  carry  10%  lead,  and  to  be 
treated  at  a  works  having  an  output  of  400  tons  of  ore  daily.  The 
cost  per  ton  of  materials  and  of  ore  are : 


OF    THE    COMMON    METALS.  335 

Materials 

of 

charge.  Ore. 

Labor                                                                              $1.10  $1.40 

General  expense,  assaying  and  management            0.20  0.27 

Fuel  for  power                                                            0.07  o.io 

Coke  (15%  of  the  charge)                                         0.97  1.36 

Limestone  0.37  tons  @  $i  0.37 

Interest,  improvement  fund  and  repairs                 0.26  0.50 

Iron  ore  o.i  ton  @  $5                                                     ...  0.50 


Total  $2.60  $4.50 

We  thus  have  the  cost  per  ton  of  ore  $4.50  and  per  ton  of  ma- 
terials $2.60.    The  above  may  be  stated  as  follows : 

1.4  tons  materials  (i  ton  ore,  0.3  ton  limerock,  and  o.i  ton  iron 

ore)  at  $2.60  per  ton  for  smelting  $3-63 

Cost  of  fluxes  0.37  ton  limerock  @  $i  0.37 

o.io  ton  iron  ore  @  $5  0.50 


$4-50 

which  corresponds  to  the  first  calculation.  In  case  the  ore  con- 
tains sulphur  in  quantity  to  need  roasting,  $2  should  be  added  to 
this  actual  cost. 

The  application  of  this  method  of  figuring  costs  may  be  illus- 
trated for  a  silicious  ore.  Let  us  suppose  by  calculation  we  have 
found  that  for  one  ton  of  such  ore  there  is  needed  1,800  Ib.  iron 
ore  and  1,450  Ib.  of  limestone.  Then  we  have: 

5,250  Ib.  or  2.625  tons  of  materials  to  be  smelted  at  $2.60  $6.82 
1, 800  "  "  0.9  "  of  iron  ore  @  $5  4-5° 

1,450  "  "    0.725    "     of  limestone  @  $i  0.72 


$12.04 

This  $12.04,  therefore,  represents  the  actual  cost  of  reducing  a 
ton  of  silicious  ore,  where  fluxes  are  paid  for  outright,  when  run 
in  a  furnace  having  suitable  lead-bearing  ores.  In  general,  how- 
ever, the  lacking  iron  for  the  charge  is  made  up,  at  least  in  part, 


336  THE    METALLURGY 

by  the  iron  excess  of  roasted  iron  sulphides  or  other  irony  ores. 
Such  an  ore  would  make  much  slag,  which  would  carry  away 
values  according  to  its  quantity ;  38%  of  this  charge  is  silicious 
ore,  and  the  room  thus  taken  up  by  fluxes  is  called  'displacement,' 
to  be  allowed  for  in  calculating  profits  because  of  the  smaller 
tonnage  of  paying  ore  put  through. 


PART  VIII.    ZINC 


PART  VIII.     ZINC 

105.     PROPERTIES  OF  ZINC. 

Zinc  is  a  white  and  brittle  metal  of  7.4  specific  gravity.  In 
commercial  form  it  is  called  'spelter.'  It  boils  at  920°  C,  and 
the  fumes  condense  between  415°  and  550°  C.  If  the  vapor  en- 
counters the  air  a  white  smoke  of  ZnO  forms. 

106.     ORES  OF  ZINC. 

The  common  ores  of  zinc  are : 

Sphalerite. — The  rosin-colored  sphalerite  is  called  rosin-blende. 
When  dark  in  color  (due  to  the  presence  of  a  little  iron)  it  is 
termed  black  jack. 

Willemite  is  a  zinc  silicate,  which  receives  the  common  name 
of  calamine.  At  a  high  temperature  in  the  retort  it  is  completely 
reduced  by  carbon. 

Smithsonite  or  zinc  carbonate  is  decomposed  on  heating,  the 
CO2  being  driven  off.  The  name  calamine  is  also  applied  to 
this  ore. 

The  largest  market  for  zinc  ores  is  at  Joplin,  Mo.,  and  here 
ores  are  bought  on  a  basis  of  60%  contained  zinc,  and  with  less 
than  i%  iron  and  2%  lead.  This  basis  price  used  to  be  reckoned 
as  being  7  times  the  St.  Louis  quotations  of  the  price  of  spelter. 
Thus,  when  spelter  was  quoted  at  $5.20  per  hundred  pounds,  the 
price  of  zinc  ore  would  be  $36.40  per  ton  of  2,000  Ib.  This  is  an 
arbitrary  calculation,  and  when  zinc  ores  are  in  much  demand 
they  command  a  better  price  than  this.  For  zinc  ores,  which 
fall  short  of  the  above-named  requirements,  deductions  are 
made  as  follows :  For  all  zinc  less  than  60%  deduct  $i  per  unit 
of  i%.  All  lead  in  excess  of  2%  takes  a  penalty  of  60  cents 
per  unit.  For  example  let  us  take  the  case  of  an  ore  containing 
58.5%  Zn  and  3.3%  Pb  and  0.7%  Fe;  thus  we  have 


34O  THE    METALLURGY 

Basis  price  on  zinc  ore  of  60  Zn  '  $36.40 

Deductions   for  zinc  60 — 58. 5  =  1. 5 X$i  =$1.50 
Deductions    for   lead   3.3 — 2=1.3X0.60=  0.78         2.28 


Net  price  paid   =$34.12 

107.     METALLURGY  OF  Zixc. 

Roasting. — For  the  successful  roasting  of  zinc  ores  a  higher 
temperature  is  needed  than  for  the  roasting  of  lead  or  copper 
ores.  Zinc  sulphate  is  formed  in  the  roasting,  which  is  only 
decomposed  at  the  highest  heat  of  the  roaster.  We  have,  in 
the  roasting  of  blende,  the  following  reactions: 

ZnS  +  3O  =  ZnO  +  SCX 
43000  86400      71000  =  -)- 114400 

2  ZnS  +  70  =  ZnSO4  +  SO2  +  ZnO 
43000  230000     71000  =  258000 

and  at  the  highest  heat  of  the  roaster 

ZnSO4  =  ZnO  +  SO3 

230000     86400     71000  =  —  72600,  an  endothermic  reaction. 

In  roasting  blende,  however,  it  is  difficult  to  reduce  the  sul- 
phur contents  to  below  i%  and  this  means  a  loss  of  2%  of  zinc. 
ZnS  contains  33%  S  to  67%  Zn. 

Reactions  in  the  retort. — When  the  zinc  ore  is  charged  into 
the  retort,  together  with  40  to  50%  of  its  weight  of  coal,  the  fol- 
lowing reactions  occur: 

Undecomposed  ZnSO4  is  reduced  to  ZnS 
Fe2O3     "  "    FeO 

ZnO        "         "         "    Zn 
The  zinc  furnace. 

The  only  method  of  recovering  zinc  from  its  ores  is  by  dis- 
tillation performed  in  retorts,  the  zinc  ore  at  white  heat  giving 
off  its  vapor  to  be  condensed  into  molten  form  in  the  cooler 
condenser. 

Fig.  140  and  141  represent,  in  transverse  section  and  in  front 
elevation,  a  Western  zinc  furnace  containing  256  retorts.  As 


OF    THE    COMMON    METALS. 


341 


342  THE    METALLURGY 

shown  in  140  there  is  a  middle  wall  which  supports  the  rear  end 
of  the  retorts.  The  front  end  of  these  retorts  passes  through 
the  front  walls  of  the  furnace.  There  are  four  deep  fireplaces 
extending  beneath  the  retorts  and  fired  by  the  four  end-doors 
with  a  soft  coal  fire,  four  feet  deep,  which  makes  them  in  reality 
gas  producers.  The  flame  from  the  fire  fills  the  whole  space 
around  these  retorts,  then  passing  off  by  outlet  ports  in  the  roof 
of  the  furnace,  and  finally,  by  flues,  to  the  two  stacks,  one  at 
each  end  of  the  'block',  as  it  is  called.  In  the  figures  the  three 


FIG.  142.     ZINC-SMELTING  RETORTS. 

lower  rows  of  the  retorts  are  seen  provided  with  condensers  in 
which  the  zinc  collects.  Beneath  the  grates  is  ample  room  for 
a  man  to  walk,  and  with  a  long  cutter  bar,  to  'grate'  or  clean 
the  fires,  removing  clinkers  and  ashes.  The  working  floor,  2  ft. 
below  the  lowest  retorts,  is  at  the  ground  level.  The  block 
proper  which  contains  the  retorts  is  14  ft.  high,  24  ft.  long  and 
12  ft.  from  face  to  face.  The  single  retort  with  its  condenser 
and  the  surrounding  brickwork  of  the  front  and  back  walls 
of  the  furnace  are  shown  in  Fig.  142.  The  retort  is  4  ft.  long 
and  8.5  in.  internal  diameter. 

Retorts  to  withstand  the  high  heat  and  corrosive  action  of 
the  charge  have  to  be  made  of  the  most  compact  and  durable 
of  fire-clay.  The  material  is  composed  of  a  mixture  of  chamotte 
'cement'  or  'grog',  being  burned  fire-clay  or  old  broken  retorts, 
brick  and  tile  coarsely  ground  to  3  or  4  mesh  and  mixed  with  an 
equal  amount  of  raw  clay.  This  is  mixed  in  a  pug-mill  with 
about  10%  of  water  to  form  a  stiff  mud  or  'adobe/  and  is  al- 
lowed to  stand  for  two  to  four  weeks  covered  with  wet  sacking 
to  develop  the  plasticity  of  the  clay.  It  is  then  again  molded, 
and  finally  made  into  retorts  by  a  hydraulic  retort-making  ma- 


OF    THE    COMMON    METALS.  343 

chine  under  a  pressure  of  3,000  Ib.  per  square  inch.  A  machine  of 
this  kind  will  make  8  retorts  per  hour.  In  Kansas  a  retort  costs 
5oc.  A  clay  commonly  used  for  retorts  contains  30%  A12O3 ; 
50%  SiO2;  15.0%  combined  water;  and  5%  of  bases. 

ZnO  is  reduced  to  Zn  and  is  volatilized,  and  later  condensed 
to  spelter.  We  have 

ZnO  +  C  •=  Zn  +  CO 

86400  29000  =  —  57400 

the  CO  escaping  at  the  end  of  the  condenser  and  there  burning, 
its  characteristic  flame,  however,  being  masked  by  a  small 
amount  of  zinc  vapor  escaping  and  burning  at  the  same  time. 
The  reaction,  as  may  be  seen,  is  quite  endothermic. 

PbO  is  reduced  to  metallic  lead  which,  in  the  small  quantity 
present  remains  in  the  retort.  CdO  is  reduced  to  Cd  and  being, 
like  the  zinc,  volatilized,  comes  away  with  it,  alloying  the  zinc 
sometimes  to  the  extent  of  0.5%,  but  most  frequently  only  to  a 
trace.  It  has  not  been  found  to  be  injurious. 

Silver  and  gold  remain  in  the  residues.  In  Missouri,  where 
there  is  little  or  none,  the  residues  are  thrown  away,  but  Colo- 
rado ores,  carrying  considerable  silver  and  a  little  gold,  yield 
residues  well  suited  to  further  treatment  in  the  blast-furnace, 
for  the  extraction  of  these  precious  metals. 


1 

N^ 

y/      .T    * 

-  —  ',-  :  A. 

FIG.   143.     CHARGE  SCOOP. 

Retorting. — The  ore,  together  with  40%  of  fine  coal,  is  mixed 
thoroughly  in  a  horizontal  pug-mill,  or  upon  the  floor  of  the 
retort  house  with  shovels.  The  charge  for  a  retort  amounts  to 
say  60  Ib.  of  the  ore  and  40  Ib.  of  coal,  and  is  skilfully  and  rapidly 
thrown  into  the  retort  with  a  special  scoop  shovel  (Fig.  143)  fill- 
ing it  clear  to  the  back. 

An  iron  rod  is  now  run  along  the  top  of  the  charge  to  form 
a  channel  for  the  escape  of  gases,  and  the  condenser  is  adjusted 


344  THE    METALLURGY 

in  position  and  luted  with  a  loamy  clay.  To  make  the  joint 
tight  a  crescent-shaped  'stamper'  is  used,  by  which  the  clay  is 
compressed  around  the  condenser.  The  stamper  (Fig.  144)  is 
of  the  form  herewith  shown  and  before  using,  is  heated  to  a  low 
red  heat  in  one  of  the  lower  rows  of  retorts,  called  cannons. 
These  cannons  are  not  used  for  retorting,  being  left  there  to 
modify  the  direct  heat  from  the  fierce  fire  below. 


FIG.  144.     STAMPER. 

A  charge  of  60  Ib.  of  60%  zinc  ore,  if  dead  roasted,  would  con- 
tain of  zinc  oxide  45  Ib.  or  of  zinc  36  Ib.  and  would  yield  at  the 
best  95%  of  the  contained  zinc.    The  residues  from  a  zinc  charge 
of  100  Ib.  (60  Ib.  zinc  and  40  Ib.  coal)  would  consist  of 
ZnO  and  ZnS,  the  former  still  persistently  re- 
maining in  the  residues,  and  the  latter  being 

unacted  on  by  the  coal   2.5  Ib. 

Gangue  and  ash  from  both  ore  and  coal 25.0  Ib. 

Unburned    coal    (two-thirds    having    been    con- 
sumed in  the  reduction  operation 13.0  Ib. 


Making  a  total  of    40.5  Ib. 

or  a  decrease  in  weight  of  the  original  100  Ib.  of  59.5%. 

To  continue:  The  retorts  are  fired  on  and  brought  up  to  a 
white  heat  of  1,300°  C.  As  the  charge  becomes  heated,  moisture, 
volatile  gases  (as  the  result  of  the  distillation  of  the  coal)  are 
given  off  and  the  reactions,  already  specified,  are  begun  be- 
tween the  ore  and  carbon.  The  zinc  vapor,  as  it  distils,  enters 
the  cooler  condenser  and  condenses  into  a  liquid  form.  The  end 
of  the  condenser  is  loosely  plastered  over  with  a  little  of  the 
charge  mixture,  and  from  time  to  time,  the  accumulating  zinc 
is  removed  with  a  ladle,  the  aperture  of  the  condenser  being 
at  once  plastered  over  again. .  Toward  the  end  of  the  distilla- 
tion firing  is  the  most  vigorous  in  order  to  remove  the  last 


OF    THE    COMMON    METALS.  345 

traces  of  metal  from  the  charge.  As  the  charge  is  put  in  early 
every  morning,  the  ore  is  nearly  24  hours  in  the  retort. 

Replacing  retorts. — The  average  life  of  a  retort  is  40  charges ; 
some  break  after  a  charge  or  two,  and  some  have  a  long 
life.  Where  the  ore  contains  iron  it  has  a  corrosive  action  on 
the  retort,  eventually  eating  it  through.  Again,  a  retort  may 
crack,  when  it  must  be  removed  and  replaced  by  a  fresh  one. 
Any  such  accident  is  indicated  by  the  failure  of  the  slight  flame 
or  smoke,  which  generally  issues  from  the  condenser,  there  being 
on  the  contrary  an  inward  current  due  to  the  chimney  draft  suck- 
ing the  air  through  the  crack  of  the  broken  retort.  Retorts  are 
not  set  in  place  when  cold,  but  are  heated  in  a  small  coal-fired 
annealing-furnace,  are  brought  to  the  bank  of  the  retorts  red-hot, 
and  at  once  put  in  place.  Room  has  first  to  be  made  by  breaking 
out  the  old  retort  and  its  surrounding  brick  of  the  furnace  front. 

Manufacture  of  retorts. — The  success  of  the  retorting  oper- 
ation depends  upon  the  durability  of  the  retorts,  and  this,  again, 
upon  their  composition  and  manufa'cture.  Since  the  bases  of 
the  charge  attack  the  retorts,  a  silicious  mixture  would  be  more 
corroded  than  a  neutral  one,  and  hence,  the  endeavor  would 
be  to  increase  the  percentage  of  clay.  On  the  other  hand,  much 
clay  would  make  the  retort  more  fusible  and  more  liable  to 
shrink  and  crack. 

Costs  of  retorting. — Bituminous  coal  is  used  to  an  amount 
varying  from  2  to  4  tons  per  ton  of  ore,  according  to  the  char- 
acter of  the  coal. 

Herewith  we  give  the  cost  of  smelting  blende  concentrate  per 
ton  of  raw  ore,  using  coal  for  fuel.  This  includes  the  cost  of 
roasting  in  the  hand-stirred  reverberatory  roaster. 

Labor  (except  for  repairs  and  renewals) $6.62 

Fuel,  3  tons  at  $0.75 2.25 

Fine  coal  or  coke  for  reduction  y2  ton  per  ton  of 

ore  at  $0.84 0.47 

Clay  for  retorts  o.i  ton  per  ton  of  ore  at  $2.60.  .  .  .  0.26 
Repairs,  renewals  and  sundry  supplies  also  putting 

repaired  furnaces  in  operation   (heating  up)  ...   0.75 

$10.35 


346  THE    METALLURGY 

Recovery  or  extraction. — This  depends  upon  the  grade  of  the 
ore,  and  while  on  an  ore  (such  as  is  worked  in  Germany)  con- 
taining 25%  zinc  the  recovery  would  be  75  to  85%  ;  in  the  Joplin, 
Mo.,  district,  on  basis  60%  Zn,  an  extraction  of  from  85  to  95% 
has  been  attained. 

Losses  of  zinc. — They  occur  both  in  roasting  and  in  retorting. 
In  roasting  there  is  a  volatilization  and  a  dust-loss  and  also  a 
loss  due  to  the  undecomposed  ZnS,  one  per  cent  of  sulphur  lock- 
ing up  2%  of  zinc.  In  retorting  there  is  an  evident  loss  at 
the  mouth  of  the  condenser  due  to  the  escaping  zinc  flame. 
Then,  since  the  walls  of  the  clay  retort  are  porous,  the  zinc  vapor 
filters  through  under  the  suction  of  the  furnace  draft.  This  is 
lessened,  however,  by  dampering  at  the  stack  and  by  making 
a  forced  undergrate  pressure.  Zinc  is  also  to  be  found  in  the 
residues,  especially  near  the  mouth  of  the  retort,  where  the  heat 
is  less.  It  may  amount  to  from  i  to  4%.  These  various  losses 
account  for  the  extractions  already  given. 

Impurities  in  the  ore. 

Iron  exists  in  many  ores.  In  the  Joplin  district  anything  over 
2%  iron  is  considered  a  detriment  to  the  ore,  lessening  the  price 
in  consequence.  Iron  has  a  corrosive  action  on  retorts,  uniting 
itself  to  the  silica  of  the  retort  and  slagging  and  softening  it 
and  eventually  eating  through.  A  remedy  for  this  is  the  use 
of  retorts  lower  in  silica  and  higher  in  alumina.  Nevertheless, 
at  the  Yieille  Montaigne  works  in  Belgium,  ores  containing  as 
high  as  20%  Fe  are  worked.  At  Pueblo,  Colo.,  Leadville  ores, 
high  in  iron,  are  also  treated,  the  iron  in  this  case  being  a  chem- 
ical constituent  of  the  blende. 

At  both  Yieille  Montaigne  and  at  Pueblo,  ores  containing 
lead  are  treated.  The  lead  becomes  reduced,  and  either  remains 
mixed  with  residue,  or  accumulates  near  the  mouth  of  the 
retort.  Since,  at  the  high  heat  of  the  retort,  it  is  somewhat 
volatilized,  it  contaminates  the  zinc,  and  must  be  later  removed 
by  a  refining  operation. 

Sadtler  process. — To  counteract  the  corrosive  action  of  bases, 
Prof.  B.  Sadtler  has  patented  a  process  which  consists  in  lining 
the  retort  writh  a  basic  lining  0.125  in.  thick.  The  lining  is  sufn- 


OF   THE    COMMON    METALS.  347 

ciently  effective  to  double  the  life  of  the  retort,  so  that  it  fails 
by  cracking  rather  than  by  corrosion.  The  process  is  in  use 
at  the  works  of  the  Cherokee-Lanyon  Spelter  Co.  near  Joplin, 
where  Colorado  ores  containing  considerable  iron  and  lead  are 
successfully  treated. 


PART  IX.    REFINING 


PART   IX.     REFINING. 

108.     REFINING  OF  METALS. 

General  phenomena  underlying  the  refining  'of  metals. — 
It  will  be  noticed  that  in  the  separation  of  metals  from  one 
another  and  from  the  contained  impurities  it  is  extremely  diffi- 
cult to  complete  the  separations.  Thus,  in  spite  of  the  pains 
taken  to  so  do,  the  commercial  metals  contain  traces  of  these 
impurities.  Metals  thus  prepared  are  therefore  graded  accord- 
ing to  quality  and  take  different  prices  in  market  according  to 
those  grades.  Thus,  Lake  copper  commands  the  highest  price 
because  of  its  purity  and  toughness,  while  casting  copper  is  re- 
served for  making  copper  castings  or  for  brass-making. 

This  is  exemplified  in  chemistry  in  the  precipitation  of  BaSO4 
from  a  BaQ2  solution,  by  means  of  H2SO4.  For  all  that  the 
reaction  is  so  perfect,  we  find  a  little  BaSO4  still  remaining  in 
solution.  On  the  other  hand,  the  precipitate  is  not  perfectly 
freed  from  BaCl2,  even  after  thorough  washing.  In  silver-lead 
blast-furnace  smelting  the  slag,  no  matter  how  perfectly  sep- 
arated, still  contains  0.2  to  0.3  oz.  silver  and  0.3  to  0.4%  of 
lead.  In  copper  refining  arsenic,  antimony  and  bismuth,  occur- 
ring in  the  crude  or  blister  copper,  is  retained  in  traces,  even 
after  refining,  and,  where  the  original  blister  copper  is  impure, 
no  high-grade  product  can  be  expected.  In  the  separation  and 
deposition  of  copper  by  electrolysis  the  same  conditions  hold. 
At  low  current  densities,  even  in  presence  of  impurities  in  solu- 
tion, the  copper  is  of  the  highest  grade ;  still  traces  of  the  im- 
purities do  find  their  way  into  the  cathode  copper,  less  so,  how- 
ever, than  in  any  other  system  of  refining. 


352  THE    METALLURGY 

109.     PARKE'S  PROCESS  FOR  THE  REFINING  OF  BASE-BULLION. 

To  get  a  clear  idea  of  refining  of  base-bullion  or  work-lead  it 
is  necessary  to  know  what  is  its  composition  of  which  we  give 
an  unusually  base  example  as  follows : 

Lead    96.59% 

Impurities  :      Cu 0.82 

As 0.38 

Sb 0.71 

Fe 0.02 

S 0.14 

2.07 

Precious  metals  Ag..i.o7       (322.0     oz.  per  ton) 

Au.. 0.0007  (     °-2°  oz-  Per  ton)    T-°7 


9973 

In  refining,  the  problem  is  to  soften  the  base  bullion  by  remov- 
ing the  impurities  and  from  the  softened  lead  taking  out  the 
precious  metals.  In  studying  the  process  the  student  should  often 
refer  to  Fig.  157. 

Softening. — This    is    performed    in    a    reverberatory     furnace 

(Fig.  145)- 

The  hearth  of  the  furnace  is  surrounded  by  a  water-jacket  a 
to  assist  the  brickwork  in  resisting  the  action  of  the  molten 
litharge,  and  is  heated  by  a  fire-box  f.  The  base  bullion,  in 
charges  of  say  30  tons,  is  charged  into  the  furnace  by  means  of 
a  long-handled  paddle  or  peel.  The  charge  is  melted  down,  fill- 
ing the  furnace  with  lead,  the  dross  rising  to  the  top.  After 
two  hours  the  dross,  which  has  had  the  lead  sweated  out  of  it,  is 
skimmed  off  with  a  long-handled  perforated  paddle.  This  dross, 
residue  or  skimming,  called  the  'copper  skim'  contains  the  iron, 
sulphur  and  especially  much  of  the  copper  of  the  base-bullion. 
The  heat  is  now  raised  to  a  bright  red,  and,  with  an  oxidizing 
flame  sweeping  over  it,  litharge  keeps  forming  and  the  arsenic 
and  antimony  oxidize  and  enter  the  litharge  slag.  This  action 
is  kept  up  for  12  hours,  or  until,  by  inspecting  a  sample  taken 


OF   THE    COMMON    METALS. 


353 


from  the  furnace,  the  lead  shows  itself  free  from  arsenic  or  anti- 
mony. The  furnace  is  allowed  to  cool,  and  when  the  litharge 
crust  is  set  or  become  solid,  it  is  skimmed  off.  For  impure  bul- 
lion, this  operation  of  heating  up,  oxidizing  and  again  cooling 
may  have  to  be  repeated  in  order  to  remove  the  last  traces  of 
arsenic  and  antimony. 


FIG.    145.     SOFTENING   FURNACE. 


The  softened  lead  is  tapped  into  a  hemispherical  de-silverizing 
kettle,  8  ft.  diam.,  capable  of  holding  readily  30  tons  of  lead, 
and  set  in  brickwork  with  a  fire-box  below.  It  is  heated  well 
above  the  melting  point  of  zinc,  and  about  1.2%  zinc  (but  vary- 
ing with  the  richness  in  silver)  added  to  the  lead.  Using  the 
Howard  apparatus  for  de-silverizing  the  process  is  then  as  fol- 


354 


THE    METALLURGY 


lows.  Fig.  146  represents  the  kettle  and  the  apparatus  used  for 
mixing  the  molten  zinc  intimately  with  the  lead.  The  machine 
is  lowered  into  the  kettle  in  the  position  shown  at  a  and  the 
propeller  b  set  in  motion  by  a  steam-driven  mechanism,  so  as 
to  produce  a  downward  flow  in  the  cylinder  a,  lead  at  the  same 
time  flowing  in  from  above.  Thus  a  thorough  mixing  of  the 
contents  of  the  kettle  is  assured,  the  mixing  being  kept  up  for 
about  10  minutes.  The  stirrer  is  then  raised  and  moved  to  one 


FIG.    146.     HOWARD    MIXER. 

side,  traveling  on  the  crawl  h.  The  contents  of  the  kettle  are  al- 
lowed to  cool  down  for  about  two  hours,  by  which  time  the 
'zinc  crust,'  a  lead-zinc-silver  alloy,  lighter  and  more  infusible 
than  the  lead,  rises  to  the  surface.  In  other  words,  the  zinc, 
having  a  greater  affinity  for  gold,  silver  and  copper  than  lead, 
removes  it  from  the  latter  as  a  crust  or  alloy.  Such  a  crust  may 
contain  JQ%  Pb,  15%  Zn  and  8%  Ag. 

Fig.  147  represents  the  press  by  which  the  solidified  crust  is  re- 
moved from  the  surface  of  the  molten  lead.  It  consists  of  a  cyl- 
inder o  which  is  partly  lowered  into  the  lead.  The  follower  c 


OF    THE    COMMON    METALS. 


355 


is  raised,  and  the  zinc  crust  skimmed  into  it  by  means  of  a  per- 
forated flat  skimmer.     Fig.  148.     When  full,  the  press  is  raised 


FIG.  147.     HOWARD  PRESS. 


and  the  plunger  brought  down  in  the  cylinder,  squeezing  out  the 
surplus  lead,  which  escapes  through  the  perforated  bottom  b, 
The  press  is  moved  to  one  side,  the  hinged  bottom  b  is  dropped 


356 


THE    METALLURGY 


and  the  zinc  crust  pushed  out,  falling  upon  the  cast-iron  floor- 
plates,  where  it  is  broken  up  in  pieces  of  fist-size.  The  operation 
of  skimming  and  pressing  is  then  repeated  until  the  lead  is  well 
skimmed.  This  crust  may  amount  to  3,000  Ib.  and  contain  90% 
of  the  silver  originally  in  the  softened  lead.  This  first  zincing 
has  not  removed  all  the  silver  and  the  oper- 
ation must  be  gone  through  with  once  or 
twice  more  before  the  contents  are  brought 
down  to  a  fraction  of  an  ounce  per  ton. 
The  second  and  later  crusts,  thus  obtained, 
contain  comparatively  little  silver  and  are 
returned  to  a  following  charge  where  their 
contained  zinc  becomes  effective  in  remov- 
ing other  silver. 

The  de-silverized  lead  in  the  kettle  still 
retains  about  half  of  the  zinc  originally  put 
in  it,  which  must  be  removed.  This  is  done 
by  tapping  it  to  a  calcining  furnace,  a  re- 
verberatory  like  the  softening  furnace 
already  described.  Here  the  charge  is 
brought  to  a  bright-red  heat,  which  results 
in  volatilizing  and  burning  off  the  zinc, 
some  litharge  being,  at  the  same  time, 
formed  by  the  scorification  of  the  lead. 

Finally  the  lead  goes  to  a  market  kettle 
similar  to  the  de-silverizing  kettle.  This  is 
the  reservoir  from  which  it  is  drawn  to 
be  cast  into  molds.  The  molds,  standing  in 
a  semi-circle,  hold  100  Ib.  of  lead  each,  and 
FIG  147.  HOWARD  PRESS  are  mounted  on  two  wheels  by  which, 
when  full  of  lead,  they  are  transferred  to 

an  adjoining  place  to  be  tilted  out  of  the  molds.  The  lead  is 
withdrawn  from  the  kettle  by  means  of  a  syphon  and  descends 
into  a  small  cast-iron  pot  into  which  is  screwed  the  2-in.  pipe 
which  delivers  the  lead  to  the  molds,  this  pipe  being  quickly 
moved  from  mold  to  mold  as  each  is  filled.  These  pigs  or  bars 
are  the  de-silverized  lead  of  commerce. 

Drv  steam  is  sometimes  blown  into  the  lead  in  the  market- 


OF    THE    COMMON    METALS.  357 

kettle  to  refine  it,  the  steam  being  introduced  by  means  of  a  pipe 
inserted  deep  beneath  the  surface.  In  this  way  the  last  traces  of 
impurities  are  removed.  It  is  softer  than  ordinary  de-silverized 
lead  and  more  easily  corroded  by  acetic  acid  in  the  process  of 
making  white  lead  and  accordingly  it  is  called  'corroding  lead/ 

Referring  to  Fig.  157,  let  us  see  what  becomes  of  the 
crust  or  skimmings,  the  result  of  the  first  zincing.  The 
material,  in  lumps  and  containing  12^  to  15%  zinc,  is  charged 
into  the  bottle-shaped  retorts  /  (Fig.  150)  each  capable  of  hold- 
ing 1,200  Ib.  of  the  crust.  Here  it  is  heated  to  a  yellow  heat 


FIG.    148.      SKIMMER. 


FIG.  149.     STIRRING  PADDLE. 

(i,ooo°C),  the  retort  being  surrounded  by  a  coke  fire,  the 
products  of  combustion  passing  off  by  the  outlet-port  seen  to  the 
left.  A  condenser,  made  by  cutting  off  the  end  of  an  old  retort, 
as  shown  in  c,  serves  to  collect  the  zinc  vapors  arising  from 
the  charge,  the  condensed  zinc  being  drawn  off  into  molds  X 
through  a  one-inch  hole  bored  at  the  bottom  corner  of  the  con- 
denser. When  distillation  is  complete,  the  condenser  and  its 
supporting  truck  are  removed,  and  the  furnace  is  tilted,  the  re- 
maining 'rich  lead'  pouring  out  into  molds  like  those  used  for 
market-lead. 

To  remove  the  lead  from  this  alloy  the  English  cupelling  fur- 
nace shown  in  Fig.  152  and  153  is  used.  The  principle  of  its 
action  is  much  the  same  as  in  cupelling  for  assay,  except  that 
the  litharge  instead  of  merely  saturating  the  cupel,  flows  off 
as  it  is  formed.  Fig.  153  is  a  sectional  elevation,  showing  a 


358 


THE    METALLURGY 


fire-box  (a)  where  a  long-flaming  coal  is  burned,  the  products  of 
combustion  passing  to  the  chimney  under  the  floor  by  the  down- 
flue  marked  b.  In  so  doing  the  flame  plays  over  a  movable 
hearth  called  a  test,  not  shown  in  Fig.  153,^^  shown  mounted 


FIG.  150.     FABER  DU  FAUR  RETORT. 

on  a  carriage  in  Fig.  152.  This  test  is  hollowed  out  like  a  cupel 
sufficiently  to  hold  a  shallow  bath  of  molten  lead  to  the  depth 
of  3  in.  Fig.  155  and  156  show  two  views  of  the  oval  test  and 
its  supporting  truck,  Fig.  156,  being  an  inverted  view  of  the 
truck  and  test.  On  the  elevation  (Fig.  152)  will  be  seen  a  pipe 


OF   THE    COMMON    METALS. 


359 


which  brings  in  air  to  a  tuyere  set  into  the  back  opening  just 
above  the  hearth.     This   plays   upon   and   oxidizes   the  lead   to 


FIG.  152.     ENGLISH  CUPELLING  FURNACE. 

Section   C.D. 


A- 


Section    A. 3. 


lofl 


FIG.  153.     ENGLISH  CUPELLING  FURNACE. 

litharge,  which  as  it  forms  runs  off  by  a  shallow  groove  or  gutter 
made  at  the  front  edge  of  the  test.  To  keep  the  test  full  rich  lead 
is  fed  in  at  the  rate  of  one  to  two  tons  daily  until  the  bath 


360 


THE    METALLURGY 


has  become  rich  in  silver,  when  the  feeding  is  stopped,  the  re- 
maining lead  being  oxidized  off  leaving  the  molten  silver,  which 
is  tapped  into  molds  each  holding  upward  of  1,000  oz.  This 
silver  is  then  subjected  to  the  acid-parting  operation  described  in 
the  chapter  on  refining. 

The  first  and  third  skimmings  (if  any)   of  the  softening  fur- 
nace, as  well  as  the  silver-bearing  litharge  from  the  cupelling 


FIG.    155.     TEST  FOR   ENGLISH    CUPELLING  FURNACE. 


* 


FIG.  156.    TEST  FOR  ENGLISH  CUPELLING  FURNACE  (INVERTED  VIEW). 


furnace  are  returned  to  the  silver-lead  blast-furnace  where  the 
sulphur  of  the  charge  unites  itself  with  the  copper,  especially 
that  of  the  first  softening-furnace  skimming,  removing  it  as 
matte ;  while  the  rest,  reduced  to  lead,  again  returns  to  the  re- 
finery. 

The    second    softening-furnace    skimming    goes    to    a    small 
reverberatory,      in      shape      like      the      softening-furnace,      and 


OF   THE    COMMON    METALS. 


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362  THE    METALLURGY 

called  a  precipitating  furnace.  Here  it  is  melted  down  at  a  good 
heat  into  a  slag,  some  charcoal  added  and  stirred  in  so  as  to 
precipitate  or  reduce  a  portion  of  the  lead.  This,  falling  to  the 
bottom  of  the  bath,  accumulates  there,  while  the  supernatant 
slag,  containing  the  antimony,  is  tapped  off.  The  precipitated 
lead  bullion  later  goes  to  the  refinery  to  be  softened  and  de- 
silverized. The  antimony  slag  (nearly  free  from  silver)  when 
a  sufficient  quantity  has  accumulated,  is  melted  in  a  small  blast- 
furnace, reducing  an  antimonial  lead  which  is  sold  to  type 
founders,  while  the  slag  produced  is  thrown  away. 

no.     VARIATIONS  ix  METHODS  OF  LEAD  REFINING. 

Variations  on  the  regular  method,  already  given,  are  as  fol- 
lows : 

Instead  of  removing  the  gold  and  silver  together  in  the  first 
crust,  it  is  possible  to  take  out  the  gold  (with  some  of  the  silver) 
by  a  separate  zincing  in  which  a  moderate  amount  of  zinc  is 
used.  This  is  due  to  the  fact  that  zinc  has  a  greater  affinity  for 
gold  than  for  silver.  Since  zinc  has  even  a  greater  affinity  for 
copper,  the  little  still  remaining  in  the  base  bullion  is  taken  up 
also.  We  thus  get  a  crust  containing  all  the  copper,  all  the  gold 
and  some  silver,  and  from  this  crust  is  produced  a  dore  silver  bar. 
The  next  zincing  removes  practically  all  the  silver ;  and  the  crust 
yields  a  silver  bar  free  from  gold,  which  needs  no  parting,  and 
which  brings  the  full  price  of  commercial  silver. 

In  the  earlier  practice  of  lead  refining  on  clean  base-bullion, 
softening  was  performed  in  the  kettle,  the  metal  was  heated  to 
just  melting  temperature,  and  the  lead  was  sweated  from  the 
copper  crust,  which  latter  was  skimmed.  The  skimmed  lead 
was  then  heated  to  a  red  heat  by  vigorous  firing,  was  stirred 
and  subjected  to  the  action  of  the  air.  Since  the  quantity  of 
arsenic  and  antimony  was  limited  they  could  be  removed  in 
this  way.  For  ordinary  base-bullion  this  method  would,  however, 
be  inadequate,  and  therefore  the  softening  furnace  is  used. 

With  regard  to  the  lead  just  de-silverized,  the  contained  zinc 
may  be  removed  in  the  de-silverizing  kettle  without  the  use  of 
a  reverberatory  furnace.  This  is  accomplished  by  heating  the 


OF   THE    COMMON    METALS.  363 

lead  in  the  kettle  to  a  high  degree  and  poling  or  rabbling  it  by 
means  of  dry  steam  or  compressed  air.  The  zinc  is  thus  in 
part  volatilized,  in  part  oxidized,  together  with  a  little  of  the 
lead,  forming  a  powdery  litharge  which  is  skimmed  off,  leaving 
market-lead. 

In  the  older  way  of  stirring  the  zinc  into  the  charge  of  the 
de-silverizing  kettle,  the  work  was  done  by  means  of  paddles. 
The  paddle  shown  at  Fig.  149  had  a  handle  about  7  ft.  long.  Two 
men,  on  opposite  sides  of  the  kettle,  worked  these,  making  strokes 
as  follows :  The  paddle-blades  were  moved  downward  to  near 
the  bottom ;  then,  by  pressing  down  on  the  handles,  and  prying 
upward,  the  molten  zinc  was  stirred  into  the  lead,  the  stirring 
taking  about  15  minutes  to  complete. 

Where  machine  methods  were  not  used,  the  first,  or  silver 
crust  from  the  de-silverizing  kettle  was  transferred  to  another 
kettle,  and  there  kept  at  a  just  visible  red-heat.  In  consequence, 
segregation  occurred  by  which  the  liquid  portion  was  separated 
and  returned  to  the  next  charge,  the  dry  crusts,  rich  in  silver, 
going  to  the  retorts. 

Besides  the  regulation  method  of  obtaining  the  copper  crust 
by  removing  it  from  the  softening  furnace,  another  effective  way 
has  been  used.  This  consists  in  charging  the  base-bullion  into  an 
inclined-bottom  reverberatory,  where  the  lead,  as  fast  as  it 
sweated  or  liquated  out,  would  drain  off  to  be  collected  in  a  small 
kettle  or  sump  outside  the  furnace.  From  here  it  was  molded 
into  bars.  The  dry  residue  was  the  copper  crust. 

in.     CHARGES  FOR  REFINING  BASE-BULLION. 

In  general,  the  metallurgist  who  operates  a  smelting  works 
has  nothing  to  do  with  the  refining  of  his  product,  which  is 
sold  to  the  refinery.  If  a  smelting  works  is  distant  from  market, 
a  month  or  six  weeks  might  elapse  before  returns  would  be 
available  from  the  refining  and  sale  of  its  product.  The  smelting 
works  may  obtain  advances  on  the  product  thus  locked  up,  by 
arrangement  with  a  local  bank.  The  bank  may  agree  to  honor 
a  draft  made  upon  it  to  the  extent  of  80%  of  the  value  of  the 
shipment  of  the  base-bullion,  when  the  draft  is  accompanied  by 


364  THE    METALLURGY 

a  signed  shipping  receipt  for  it,  as  consigned  to  their  account, 
together  with  the  smelter  certificate  of  assay.  The  refinery  pays 
the  bank  for  the  full  value  of  the  product  when  refined,  and  the 
20%  balance  due  is  then  credited  by  the  bank'  to  the  reduction 
works.  This  is  often  of  great  assistance  to  a  works,  which  thus 
procures  the  money  needed  for  operating  expenses  as  soon  as 
it  has  loaded  out  the  base-bullion.  By  arrangement  with  a  refin- 
ery, a  reduction  works  may  obtain  the  following  credits  and 
charges  for  refining  per  fine  ounce  at  the  reduction  works : 

For  silver :  the  New  York  quotation  per  fine  ounce  less  one 
cent  per  oz.  (this  deduction  is  to  meet  the  expense  of  parting). 

For  lead:  the  New  York  quotation  less  I2^c  per  100  Ib.  or 
$2.50  per  ton  of  contained  lead.  The  deduction  is  made  to  meet 
the  cost  of  freighting  the  lead  from  the  refinery  to  New  York 
City. 

For  gold :  $20  per  ounce.  The  mint  value  is  $20.67  Per  oz-> 
so  that  the  refinery  obtains  a  profit  on  the  difference. 

From  these  values  must  be  deducted  the  treatment  charges  of 
$12  per  ton,  and  the  freight  (advanced  by  the  refinery)  of,  say, 
$8  per  ton. 

Compare  these  charges  with  actual  costs  of  refining  as  per  the 
following  estimate : 

Prime  cost  of  softening  and  refining.  .  .  .$5.00  to  $6.00 

General   expense 3.00  to     3.00 

Loss  in  metals  and  incidental  expenses.  .  .    1.70  to     3.00 


$9.70  to  $12.00 

112.     ELECTROLYTIC  REFINING  OF  COPPER. 

The  blister-copper  as  it  comes  in  from  the  reduction  works 
in  the  West,  is  sampled*  by  taking  drillings  from  each  bar.  These 
drillings,  mixed  together,  represent  the  bars  of  a  given  lot.  The 
copper  is  now  sent  to  the  re-melting  furnaces,  each  capable  of 
holding  150,000  Ib.  copper.  Here  it  is  somewhat  refined  by  rab- 


*  This  description  of  a  copper  refinery  is  based  on  the  practice  at  the 
Raritan  Copper  Co.,  Perth  Amboy,  N.  J. 


OF- THE    COMMON    METALS.  365 

bling  and  poling  and  it  is  then  cast  into  anodes  by  Walker  casting- 
machines.  As  the  anodes  come  from  the  machine  they  are 
trimmed  from  any  fins  of  copper  and  racked  on  trucks,  22 
together. 

The  trucks  thus  loaded  are  moved  into  the  tank-house  where 
the  anodes  are  picked  up,  the  group  of  22  together,  and  placed  in 
the  tank.  There  are  420  tanks  in  all,  each  tank  being  in  electrical 
series  with  the  next.  The  current  goes  through  each  of  the  22 
plates  or  anodes  to  the  cathodes  opposite  and  placed  about  2  in. 
apart.  Assuming  that  the  anodes  are  2  by  3  ft.,  the  current  leav- 
ing both  faces,  we  have  a  total  area  through  which  the  current 
is  passing  of  264  sq.  ft.  with  a  density  of  current  of  15  amperes 
per  square  foot.  We  thus  have  a  total  of  3,960  amperes.  These 
plates  are  immersed  in  a  solution  or  electrolyte  of  CuSO4  in 
dilute  H,SO4,  the  CuSO4  being  15%  (4%  Cu),  and  the  sulphuric 
acid  10%  of  the  solution.  The  resistance  in  passing  through  this 
space  of  about  2  inches  to  the  cathode  is  0.25  of  a  volt.  The 
tanks  in  series  would  therefore  give  a  resistance  of  105  volts. 

For  the  cathodes,  so-called  'stripping  plates'  are  prepared  as 
follows :  In  a  portion  of  the  tanks  insoluble  anodes  of  lead  are 
used,  opposite  to  which  are  set  greased  copper  plates,  upon 
which  is  precipitated  a  layer  of  copper.  When  this  layer  be- 
comes 1/32  inch  thick,  the  greased  plates  are  removed  from  the 
tank,  and  these  coatings  or  sheets  of  copper  are  stripped  off,  the 
greased  surface  having  prevented  adhesion.  Loops  of  similar 
copper  are  riveted  to  these  plates,  which,  suspended  in  the  regular 
tanks,  form  the  beginning  of  the  cathodes. 

One  ampere  will  deposit  an  ounce  avoirdupois  of  copper  in  24 
hr.,  so  that,  at  this  plant,  each  plate  increases  in  weight  nearly 
12  Ib.  daily.  At  the  end  of  two  weeks  the  completed  cathode, 
weighing  160  Ib.,  is  removed,  and  another  one,  replacing  it,  will 
receive  what  has  been  left  of  the  anode.  Of  course,  not  all  the 
copper  is  dissolved,  since  the  part  above  the  solution  and  the 
skeleton  of  the  plate,  at  least,  must  be  there  to  transmit  the  current. 
Even  then,  shreds  and  pieces  of  copper  drop  to  the  bottom  of  the 
-tank,  and  the  fragmentary  anode  is  removed,  still  having  15%  of 
its  original  weight.  They  are  re-melted  and  re-cast  into  new 
anodes. 


366  THE  METALLURGY 

113.     COSTS  OF  ELECTROLYTIC  REFINING  OF  COPPER. 

In  the  earlier  experiences  of  refining  costs  were  high,  as  much 
as  $20  per  ton,  and  refiners  charged  $40  per  ton.  At  present  the 
cost  of  refining  98%  copper  anodes  has  been  lowered  to  $4  and 
$5,  and  contracts  have  been  closed  at  as  low  as  $7.50  per  ton  of 
anodes.  This  is  due  to  economy  of  power  and  in  handling  ma- 
terial. 

Current  density. — The  greater  the  density  of  the  electric  current 
through  the  electrodes  the  faster  the  deposition ;  and  hence  the 
smaller  stock  of  metals  which  needs  to  be  carried  for  a  given 
output.  Thus  at  Great  Falls,  Mont.,  the  current  density  is  40  am- 
peres per  square  foot  of  anode-surface,  while  at  Perth  Amboy, 
N.  J.,  the  density  is  15  to  17  amperes.  At  high  density  there  is  the 
chance  of  short-circuiting,  and  moreover,  the  cathode  plates  take 
on  more  impurities,  as  antimony  and  arsenic,  from  the  bath.  It 
takes  a  month  for  an  anode  to  dissolve ;  hence,  in  a  large  refinery, 
from  half  a  million  to  a  million  dollars  may  be  tied  up  in  copper 
and  in  the  precious  metals. 

Treatment  of  the  slime  or  anode  mud. — From  time  to  time  the 
thin  mud  which  accumulates  in*  the  bottom  of  the  tanks  is  re- 
moved by  decanting  off  the  clear  electrolyte,  and  then  washing 
the  mud  into  a  tight  car  placed  below  the  outlet-plug  of  the  tank. 
The  mud  contains  fragments  of  copper  of  all  sizes,  dropped  from 
the  anodes  when  nearly  dissolved  and  which  are  mostly  removed 
in  passing  through  a  4O-mesh  screen.  The  mud  is  transferred 
to  a  settling  tank.  From  here  it  is  drawn  to  a  pressure  tank 
and  thence  to  a  filter  press,  dried  by  steam  to  2%  moisture, 
and  broken  up.  It  is  then  ready  for  treatment  on  an  English 
cupelling-hearth.  Here  it  is  brought  in  the  furnace  upon  a  lead 
bath  into  which  it  is  melted,  the  lead  taking  up  the  silver  and 
gold.  The  impurities,  as  copper,  arsenic,  antimony  and  tellurium, 
are  scorified,  going  into  the  air  and  into  the  litharge,  which  forms 
under  the  action  of  the  air-blast  commonly  used  in  this  kind  of 
furnace.  The  lead,  having  been  cupelled  away,  the  dore  silver  left, 
is  removed  and  molded  into  bars.  The  scrap  copper,  which  was  re^ 
moved  from  the  anode-mud,  as  above  referred  to,  is  re-melted  and 
re-cast  into  anodes.  Such  scrap,  together  with  the  rect  of  the  un- 


OF    THE    COMMON    METALS.  367 

dissolved  anode,  amounts  to  15  to  20%  of  the  total.  Another 
method  of  treatment  consists  in  treating  the  anode-mud  with 
dilute  sulphuric  acid  which  will  remove  the  copper. 

Purifying  the  Electrolyte. — -The  electrolyte  commonly  contains 
16%  H2SO4  and  4%  copper.  When  the  copper  exceeds  this 
quantity  the  resistance  increases,  and  hence  the  copper  should  be 
removed  from  the  system.  The  solution  also  contains  iron,  arsenic 
and  antimony.  Antimony  may,  however,  be  kept  low  by  the  daily 
addition  of  a  small  amount  of  salt,  which  precipitates  it  as  an 
oxychloride.  To  purify  the  electrolyte  the  following  method 
is  used.  A  portion  of  the  electrolyte  is  diverted  to  tanks  having 
lead  anodes  and  using  a  strong  current.  The  impurities,  together 
with  some  copper,  are  deposited  loosely  upon  copper  cathodes. 
Every  two  months  the  accumulated  mud,  containing  40  to  60% 
copper,  is  cleaned  out  and  reduced  to  impure  bars  at  a  refining 
furnace.  The  electrolyte,  thus  purified,  is  standardized  to  the 
proper  strength  and  returned  to  the  circulation. 

Circulation  of  the  electrolyte. — To  avoid  short-circuiting  and  to 
increase  the  activity  of  deposition,  the  electrolyte  is  made  to 
flow  through  the  tanks  entering  near  the  top  at  one  end,  and 
rising  through  an  overflow  pipe  at  the  other.  After  it  has  flowed 
through  two  tanks  in  this  way,  it  enters  a  launder  which  returns 
it  to  a  collecting  or  sump  tank.  Here  it  is  brought  up  to  a  tem- 
perature of  40°  C  by  the  use  of  a  steam-coil  and  is  then  pumped 
to  the  distributing  tank  for  re-circulating. 

Capital  in  plant. — Not  only  is  there  capital  invested  in  the  build- 
ings and  machinery  of  the  plant,  but  there  has  to  be  capital  pro- 
vided for 

1 i )  The  stock  of  anodes  in  process  of  treatment. 

(2)  The  stock  on  hand  for  one  month. 

(3)  The  copper  locked  up  and  permanently  in  the  electrolyte 
or  solution. 

(4)  The  copper  locked  up  in  heavy  conductors. 

Testing  the  current. — ^Besides  the  recording  instruments  to  be 
found  on  the  switch-board  of  the  generators  in  the  power-house 
it  is  customary  to  use  others  constantly  in  testing  the  drop  of  po- 
tential between  plates.  To  do  this  a  forked  rod  is  used  which 
touches  adjoining  plates  and  which  transmits  the  current  through 


368  THE    METALLURGY 

a  volt-meter.    Any  variation  due  to  short-circuiting  is  thus  known 
and  at  once  rectified. 

114.     COPPER  REFINING  FURNACE. 

Fig.  158  represents,  in  sectional  plan  and  Fig.  159  in  sectional 
elevation,  a  copper-refining  furnace,  12  by  18  ft.  in  size,  suited  to 
the  treatment  of  blister  copper,  and  of  a  capacity  of  30,000  to 
40,000  Ib.  to  the  charge.  The  stack  or  chimney  (not  shown)  is 
close  to  the  furnace,  with  a  communicating  outlet-flue  indi- 
cated in  the  elevation.  For  the  admission  of  air  there  are  air- 
ports at  the  sides  as  seen  in  the  plan,  also  a  row  of  air-holes  over 
the  bridge.  An  arched  space  is  left  below  the  hearth  for  ventila- 
tion, and  a  pit  is  provided  next  to  the  ash  pit  where  stands  the 
fireman  when  grating  or  cleaning  the  fire.  The  bridge,  as  shown 
in  the  elevation,  is  provided  with  a  hollow  casting  called 
the  conker-plate  in  which  circulates  air  for  cooling  it.  At  the 
front  end  of  the  hearth,  next  to  the  front  door,  will  be  noticed  a 
bowl-shaped  depression  or  sump,  whence  the  metal  may  be  dipped 
or  ladled  out. 

We  may  add  that  in  recent  practice  much  larger  and  deeper 
furnaces  are  used  than  here  specified,  and  having  a  hearth  dimen- 
sion 14  by  26  ft.  and  holding  over  200,000  Ib.  of  copper.  So  large 
a  quantity  of  copper  could  hardly  be  handled  by  the  slower 
methods  of  molding,  and,  for  so  doing,  copper  casting-machines 
are  used,  capable  of  molding  50,000  Ib.  of  copper  per  hour. 

115.     COPPER  REFINING. 

Blister  copper,  whether  produced  by  the  converter,  the  rever- 
beratory  furnace,  or  by  melting  down  concentrate  (mineral)  from 
the  native  copper  ores  of  the  Lake  Superior  region,  still  con- 
tains impurities,  principally  arsenic  with  some  sulphur  and  iron, 
which  must  be  removed  by  refining. 

This  is  performed  in  the  furnace  Fig.  158.  The  charge  of 
30,000  to  40,000  Ib.  in  pigs  is  put  in  at  the  side  doors  of  the  fur- 
nace. Where  it  is  fine  enough,  as  in  nearly  pure  'mineral',  it  may 
be  dropped  through  the  roof  from  an  overhead  hopper.  In  the 


OF   THE    COMMON    METALS. 


369 


37O  THE    METALLURGY 

Lake  Superior  country  large  pieces  of'  copper,  sometimes  tons  in 
weight  and  called  'mass  copper,'  are  put  in  through  a  hatch  or 
large  opening  in  the  roof,  this  hole  being  then  covered  by  a  lid  or 
cover  of  brick  bound  in  an  iron  frame.  The  furnace  is  now  vigor- 
ously fired  and  the  charge  melted  down,  this  taking  several  hours. 
The  front  door  is  then  opened  and  the  rabbling  or  flapping  of  the 
charge  takes  place.  This  consisted  formerly  in  striking  the  sur- 
face of  the  copper  with  a  rabble  in  such  a  way  as  to  splash  it  up 
exposing  it  to  the  action  of  the  air.  The  present  practice  is,  how- 
ever, to  insert  a  o.75~in.  pipe  in  the  molten  metal  by  which  com- 
pressed air  is  driven  into  it,  agitating  and  at  the  same  time  oxidiz- 
ing it.  This  action  proceeds  to  the  stage  of  set-copper,  the  arsenic 
being  oxidized  and  driven  into  the  slag  as  well  as  volatilized, 
and  the  copper  having  absorbed  an  excess  of  oxygen.  The  slag, 
formed  during  the  operation,  is  then  skimmed  off.  The  oxygen 
must  next  be  removed  by  poling,  which  is  done  by  inserting  into 
the  front  door  spruce  or  poplar  poles,  the  butt  ends  of  the  poles 
being  forced  into  the  metal.  At  the  same  time  a  wheelbarrow-load 
of  charcoal  is  thrown  in  to  assist  reduction.  This  work  takes  an 
hour  or  two,  a  sample  of  the  copper  being  taken  from  time  to 
time  until  the  copper  has  gotten  to  "tough  pitch",  which  is  shown 
by  the  appearance  of  its  fracture.  The  contents  of  the  furnace  are 
now  ready  for  dipping  out  into  molds  by  means  of  ladles.  These 
are  either  hand  ladles,  or  so  called  "bull"-ladles  capable  of  holding 
200  Ib.  of  copper  each,  carried  by  an  overhead  trolley,  and  by  which 
the  copper  is  readily  handled.  The  molds  are  of  different  forms 
or  shapes  called  for  by  the  trade;  forming  cakes  for  rolling  into 
sheets,  ingots  for  melting  into  brass,  and  wire-bars  for  making 
into  wire.  Where  the  copper  contains  silver  enough  to  justify  so 
doing,  the  copper  for  electrolytic  refining  is  cast  into  anodes. 

116.     PARTING  GOLD-SILVER  BARS. 

Nitric  acid  parting: — This  method,  still  practised  at  the  United 
States  Mint,  Philadelphia,  is  a  very  efficient  way  of  parting,  es- 
pecially when  working  on  a  small  scale.  Commercially,  however, 
parting  with  sulphuric  acid  is  cheaper,  and  is  generally  used. 

In  any  case  the  bars,  when  containing  too  great  a  proportion  of 
gold  are  inquartated  my  melting  them  with  silver  so  as  to  bring 


OF   THE    COMMON    METALS.  371 

down  the  proportion  of  gold  to  one-fourth  of  the  whoie.  In  part- 
ing by  sulphuric  acid,  the  amount  of  copper  in  the  bars  should  be 
less  than  10%,  but  in  ritric  acid  parting  more  than  10%  is  al- 
lowable. The  melt  is  granulated  by  pouring  the  metal  into  water. 
It  is  now  treated  with  strong  nitric  acid  of  1.20  sp.gr.  in  porce- 
lain, glass,  or  platinum  vessels.  When  action  of  the  acid 
has  ceased,  the  solution  is  decanted  and  fresh  acid  added  which 
dissolves  the  remaining  traces  of  silver.  This  is  again  decanted, 
and  the  gold  residue  washed  thoroughly  with  hot  water.  It  is 
then  withdrawn,  mixed  with  a  little  flux,  and  melted  in  a  graphite 
crucible.  From  the  nitric  acid  solution  the  silver  is  precipitated 
by  the  addition  of  common  salt.  The  silver  chloride  thus  formed 
is  washed  thoroughly,  zinc  added  in  quantity  sufficient  to  reduce 
the  silver  to  metallic  form,  and  the  zinc  chloride  is  washed  out 
from  the  precipitated  silver.  The  silver  is  pressed  into  cakes  and 
melted  down  in  crucibles  into  bars  for  sale. 

Sulphuric  acid  parting. — Since  in  this  method  of  parting  there 
must  be  less  than  10%  copper,  it  is  necessary  to  correctly  propor- 
tion a  melt,  hence  an  unlimited  number  of  mill  bars  high  in  copper 
would  be  inadmissible.  The  bars  or  ingots  of  the  metal  are  put 
into  a  cast-iron  kettle  covered  with  a  hood  and  treated  with  a 
quantity  of  sulphuric  acid  of  66°  B.  When  action  has  ceased, 
the  acid  is  allowed  to  settle  and  the  clear  supernatant  acid  is  drawn 
off  by  means  of  a  syphon  to  a  lead-lined  precipitating  tank,  while 
the  residue  in  the  tank  is  treated  six  or  seven  times  to  fresh  acid. 
In  this  way  the  silver  is  completely  dissolved,  the  clear  acid  being 
removed  after  each  treatment.  The  gold  residue  is  then  boiled 
with  water,  heated  and  agitated  by  live  steam  from  a  pipe  inserted 
in  the  water.  In  this  way  the  gold  is  sweetened.  It  is  removed 
from  the  kettle,  dried  and  melted  in  crucibles  with  a  little  borax 
for  flux.  The  acid  solution  from  the  kettle  containing  the  silver 
is  now  diluted  with  some  water,  and  the  silver  is  precipitated  by 
hanging  copper  plates  about  one  inch  in  thickness  in  the  solution, 
which  becomes  of  a  blue  color  from  the  presence  of  the  copper. 
When  precipitation  is  complete  the  clear  solution  is  decanted 
•and  the  cement  silver  at  the  bottom  of  the  tank  is  washed  with 
hot  water  to  remove  all  acid  and  the  copper  in  solution.  The 
cement  silver  is  then  removed  to  a  hydraulic  press  since  it  is  too 


3/2  THE    METALLURGY 

bulky  to  melt  directly  in  crucibles.  When  thus  compressed  it 
can  be  readily  melted  down  in  plumbago  crucibles  with  a  little 
nitre.  In  large  establishments  the  melting  down  is  performed 
in  an  English  cupelling  furnace  where  it  can  be  conveniently 
fluxed,  toughened  and  skimmed.  It  is  made  into  bars  of  about 
500  oz.  each.  On  refining  each  bar  is  marked  with  its  number, 
exact  weight  and  fineness,  and  with  the  name  of  the  refinery. 


PART  X.    COMMERCIAL 


PART    X.     COMMERCIAL. 

117.     LOCATION  OF  REDUCTION  WORKS. 

The  location  of  a  reduction  plant  will  depend  much  upon 
whether  it  is  a  mine  or  a  custom  works.  In  the  former  case  it 
has  been  usual  to  place  a  stamp-mill  as  near  the  mine  as  a  suitable 
site  and  water  supply  will  permit.  In  case  of  a  reduction  works, 
where  fuel  and  fluxes  are  to  be  brought  in  and  products  shipped 
out,  then  in  addition,  location  in  reference  to  the  railroad  must 
be  reckoned  on.  For  custom  works,  using  ores  from  different 
mines  and  localities,  a  point  central  to  them,  and  also  a  railroad 
center  must  be  considered. 

The  respective  advantages  of  the  side  hill  (terraced  site)  and 
the  level  site  have  been  much  discussed,  it  having  been  claimed 
that  the  former  permits  the  ore  to  proceed  by  gravity  from  point 
to  point  as  it  is  reduced  without  having  again  to  elevate  it.  The 
advantages  of  the  flat  site  are  as  follows : 

1.  The  first  cost  of  the  works  is  smaller,  since  grading  and 
retaining  walls  are  reduced  to  a  minimum. 

2.  The  arrangement  is  more  convenient  since  one  is  not  com- 
pelled to  place  the  different  parts  of  the  plant  in  a  certain  pre- 
destined order  to  obtain  the  fall  required.    Also,  when  one  wishes 
to  expand  the  works,  it  is  possible  to  do  so  in  any  direction. 

3.  Every  square  foot  of  the  ground  may  be  at  will  the  equiva- 
lent of  a  lower  or  upper  terrace  to  every  other,  and  hence  parts  of 
the  plant,  which  must  be  far  apart  on  a  terrace  site,  can  be  set 
side  by  side  on  a  level  one.     Ventilation  is  much  better  and  the 
plant  more  accessible  to  supervision. 

4.  Of  course  on  a  level  site  one  must  use  elevators,  but  it  is 
seldom  that  we  find  a  terrace  site  where  elevators  are  not  used. 
A  part  of  the  product  of  a  reduction  works  has  often  to  be  sent 
back  for  re-treatment.     The  cost  of  elevating  may  be  set  at  o.5c 


376  THE    METALLURGY 

per  ton  only.  Finally  we  find  the  iron  and  steel  plants,  the  largest 
reduction  works  in  the  world,  situated  on  level  sites. 

Mill-sites. — In  the  United  States  title  can  be  secured  from  the 
general  government  for  a  mill-site  for  reduction  works  upon  the 
non-mineral  unclaimed  public  lands  to  the  extent  of  five  acres, 
either  in  connection  with  a  mineral  claim  (on  the  theory  that  each 
lode  is  entitled  to  a  mill-site)  or  by  an  independent  reduction  com- 
pany, and  it  is  located  much  like  a  mineral  claim. 

A  reduction  company  must  take  care  of  its  own  tailing  and 
water,  must  not  permit  them  to  flow  upon  the  property  of  other 
persons,  and  is  responsible  for  all  such  damages.  Neither  must 
it  flow  tailing  into  a  stream  when  at  a  small  cost  it  can  be  im- 
pounded, nor  can  it  be  run  into  waters  where  it  is  liable  to  inter- 
fere with  navigation. 

A  reduction  company  may  also  take  up  land  for  a  ditch  or 
flume,  or  unappropriated  public  land,  which  cannot  be  inter- 
fered with  by  later  locators ;  but  the  owner  of  such  ditches  is 
responsible  for  damage  arising  from  their  breaks  or  overflow. 
The  same  rules  hold  in  respect  to  trails  and  roads.  In  Colorado 
all  claims  are  subject  to  the  right-of-way  of  parties  hauling  ore 
over  them ;  but  elsewhere  the  location  gives  exclusive  claim  except 
by  a  water,  electric  or  railroad  company  under  the  law  of  eminent 
domain,  and  even  then  with  fair  compensation. 

The  right  or  custom  of  dumping  across  the  valueless  land  of 
lower  mineral  claims  is  -general,  except  that  it  must  not  do  damage 
to  the  owners  below. 

118.     HANDLING  OF  MATERIALS. 

For  the  horizontal  handling  of  a  hundred  tons  or  less  of  ma- 
terial daily,  especially  when  it  is  to  be  distributed  to  various 
places,  one  and  two-wheeled  wrheelbarrows  and  two-wheeled  bug- 
gies have  been  found  economical,  elastic,  and  involving  but  little 
first  cost.  For  larger  quantities  the  same  appliances  are  used, 
as  well  as  hand-propelled  tram-cars  as  in  mine  work.  For  larger 
quantities  power-propelled  cars  are  used,  which  permit  of  being 
handled  on  up-grades,  such  as  would  be  disadvantageous  in  the 
case  of  hand-moved  appliances. 


OF    THE    COMMON    METALS.  377 

The  application  of  machinery  to  the  handling  of  materials 
has,  of  late  years,  received  great  attention,  and  has  been  rap- 
idly developed,  because  of  the  economies  in  labor  brought  about 
by  its  use. 

The  methods  may  be  divided  into: 

1 i )  Appliances  for  moving  materials  continuously,  as  belt  ele- 
vators and  conveyors. 

(2)  Appliances  for  handling  material's  intermittently,  as  light 
railways,  ropeways  and  cranes. 

( i )  Continuous-type  machines  carry  a  distributed  load,  permit- 
ting of  lighter  construction.  They  also  have  the  advantage  that 
the  delivery  is  continuous,  no  time  being  lost  in  the  operations 
of  charging  and  discharging.  Intermittent  work,  on  the  con- 
trary, if  we  increase  the  load  of  the  skip  or  bucket,  becomes  awk- 
ward and  slow  in  handling,  whereas  in  a  continuous  conveyor,  it 
is  possible,  for  example,  to  quadruple  capacity  by  increasing  the 
width  of  the  conveyor,  provided  that  the  feed  be  correspondingly 
increased.  Continuous  machines  may  be  divided  into : 

(a)  Elevators  for  vertical  lifting. 

(b)  Conveyors  for  horizontal  conveying. 

(c)  Appliances  which  perform  both  duties  at  the  same  time, 
(a)   Elevators. — The  well-known  belt-elevator  is  of  this  type, 

consisting  of  an  endless  leather  or  rubber  belt,  having  sheet- 
steel  buckets  attached  to  it  at  intervals  of  say  18  in.  Fig.  160  rep- 
resents such  an  elevator.  To  allow  for  the  stretching  of  the  belt 
the  upper  pulley  is  carried  in  take-up  boxes.  The  lower  pulley  is 
inclosed  in  a  boot,  the  ore  delivering  into  it  on  the  rising  side  of 
the  belt  at  the  left.  The  boot  has  a  hinged  drop-bottom  for  clean- 
ing out.  As  the  ore  accumulates  in  the  boot  it  is  scooped  up  by 
the  buckets,  so  that,  provided  the  feed  is  not  excessrve,  it  is  self- 
cleaning.  Fig.  161  shows  a  pressed-steel  bucket,  which  is  at- 
tached to  the  belt  by  flat-headed  elevator-bolts.  The  speed  must 
be  such  as.  to  ensure  by  centrifugal  action  the  delivery  of  the 
ore  to  the  discharge  spout  as  shown  by  the  arrow. 

With  a  belt-speed  of  275  to  300  ft.  per  min.  the  capacity  of  an 
elevator,  having  a  bucket  8  in.  wide,  will  be  7  tons  per  hour,  and, 
for  a  lo-in.  bucket,  18  tons  per  hour.  The  head  and  boot  pulleys 
are  30  in.  diam.,  and  i  in.  wider  than  the  width  of  the  bucket. 


378 


THE    METALLURGY 


FIG.  160.    VERTICAL  BELT  ELEVATOR. 


OF    THE    COMMON    METALS. 


379 


Fig.  162  is  a  single  strand  endless-chain  elevator.  The  head  and 
foot  are  carried  by  sprocket  wheels  with  teeth  spaced  to  suit  the 
chain.  In  this  case  the  take-ups  are  carried  by  the  boot.  Fig.  164 
represents  either  of  these  elevators  encased  in  wood. 


FIG.   161.     STEEL  ELEVATOR  BUCKET. 


FIG.  162.     SINGLE-STRAND 
ENDLESS-CHAIN  ELEVATOR. 


FIG.  164.     ENDLESS-CHAIN  ELEVATOR. 


THE    METALLURGY 


Conveyors. — The  spiral-worm  conveyoY  is  convenient  for  the 
delivery  of  crushed  ore  for  short  distances ;  it  thoroughly  mixes 
the  ore  in  transit.  Its  disadvantages  are  the  power  required,  the 
grinding  action  on  the  ore,  and  the  wear  and  tear  of  the  apparatus. 


FIG.  165.     SCREW  CONVEYOR.     (QUARTER  TURN.) 

Fig.  165  represents  such  a  conveyor,  delivering  ore  from  one 
box  to  another  at  right  angles,  the  ore  dropping  from  the  first 
box  upon  the  spiral  conveyor  in  the  second  one. 

Fig.  167  shows  such  a  conveyor  fitted  with  a  tripper  by  which 
the  ore  is  delivered  to  any  desired  bin.  It  will  be  noticed  in  regard 


FIG.  167.     CONVEYING  BELT  WITH  TRIPPER. 


OF    THE    COMMON    METALS.  381 

to  these  conveyors,  that  they  can  carry  ore  not  only  on  a  level 
but  at  as  steep  an  angle  as  24°.  Their  capacity  is  large;  thus  a 
12-in.  belt,  traveling  at  the  rate  of  150  to  350  ft.  per  minute,  will 
deliver  from  10  to  35  tons  per  hour.  A  24-in.  belt,  traveling  600 
ft.  per  minute,  would  have  a  maximum  capacity  of  250  tons  per 
hour  on  crushed  ore,  and  would  take  6  h.p.  for  a  conveyor  100 
ft.  long.  If  the  ore  goes  up  an  incline,  the  power  needed  is  much 
greater.  These  conveyors  are  simple  and  durable. 

Fig.  1 68  represents  an  endless  chain  'push'  conveyor,  a  series 
of  'flights'  being  attached  to  the  double  chain  by  which  the  ma- 


FIG.  168.     ENDLESS-CHAIN  PUSH  CONVEYOR. 


terial,  drawn  from  storage  bins,  is  scraped  along  a  fixed  sheet- 
iron  lined  trough,  and  even  up  an  incline,  where,  in  this  case,  it 
is  delivered  to  an  endless  chain  elevator.  The  drawback  to  these 
conveyors  is  the  wear  of  the  numerous  joints  and  of  the  trough 
and  flights,  and  the  limited  velocity  as  compared  with  the  belt- 
conveyor. 

Vibrating-trough  conveyors.  This  conveyor,  see  Fig.  169,  con- 
sists of  a  sheet-steel  flat-bottomed  trough,  say  50  ft.  long  by  2 
ft.  wide,  held  up  by  spring-legs  set  at  an  angle  and  receiving 
from  an  eccentric  a  throw  of  I  in.  at  the  rate  of  300  rev.  per 
minute. 


THE    METALLURGY 


.  These  conveyors,  in  series,  may  be  made  to  carry  to  a  distance 
of  500  ft.  Besides  the  horizontal  they  have,  owing  to  the  slope  of 
the  spring  legs,  a  vertical  movement,  as  given  in  the  diagram  (  Fig. 
170),  where  E  Et  E2  and  E3  show  the  successive  position  of  par- 
ticles of  the  ore  being  conveyed.  The  full  lines  show  the  con- 
veyor at  one  end  of  its  stroke,  the  dotted  lines  the  other,  the 
vertical  motion  being  0.125  in.  These  conveyors  are  simple, 
inexpensive  and  require  but  little  repairs.  A  24-in.  conveyor  has 


FIG.  169.    VIBRATING  TROUGH  CONVEYOR. 

a  capacity  of  20  to  25  tons  per  hour.  These  conveyors  are  well 
suited  also  for  screening  ore. 

(2)     Appliances  for  handling  materials  intermittently. 

(a)  Hoists.  Of  these,  the  most  common  about  reduction 
works  is  the  vertical  platform-elevator  which  takes  buggies,  cars 
or  wheelbarrows  from  floor  to  floor.  It  has  a  platform,  large 
enough  to  take  two  wheelbarrows,  say,  6  ft.  by  6  ft.  area.  These 
elevators  raise  a  load  of  one  ton  at  60  ft.  per  minute,  and  are 
often  run  in  balance,  though  it  is  more  certain  to  have  two,  each 
platform  being  counterweighted.  Necessarily,  time  is  lost  in 
loading  and  unloading,  so  that  one  can  figure  on  no  more  than  15 
tons  capacity  hourly. 

We  have  in  Fig.  96  the  iron  blast-furnace-hoist,  showing  the 
method  of  loading  and  of  dumping.  The  capacity  of  the  skips  is 


OF    THE    COMMON    METALS.  383 

generally  two  tons  of  ore  and  one  of  coke,  and  they  are  run  in 
balance.  It  takes  34  seconds  actual  time  for  raising,  dumping  and 
returning,  but  the  time  for  each  turn,  including  the  waits  for  load- 
ing, is  about  4  minutes,  making  trips  enough  for  a  capacity  of 
350  to  400  tons  of  pig-iron  daily. 

(b)    Conveying  appliances. 

Industrial  railroads  or  tram  tracks.  These  are  used  about 
large  reduction  works  for  delivering  materials,  not  only  for  points 
on  the  same  level,  but  also  to  move  them  up-grade,  taking  ad- 
vantage either  of  a  side-hill  or  of  trestles  in  so  doing.  Steam, 
electricity  and  compressed  air  have  all  been  used,  the  latter  being 


//////////////w/^ 

FIG.   170.     DIAGRAM   SHOWING  ACTION  OF  SWINGING  CONVEYOR. 

successful  because  of  its  simplicity,  and  because  the  short  runs 
that  are  made  permit  frequent  charging. 

Fig.  171  shows  an  electric  trolley  system  with  motor  for  hand- 
ling a  31  cu.  ft.  slag-car,  as  used  at  the  United  Verde  mine, 
Jerome,  Arizona. 

119.     ORGANIZATION  OF  A  METALLURGICAL  COMPANY. 

To  perform  metallurgical  operations  on  a  commercial  scale  re- 
quires generally  the  organization  of  a  company,  or,  if  a  mining 
company  is  already  organized,  the  addition  of  a  department  which 
will  take  care  of  such  additional  functions. 

In  the  first  case,  the  promoters  or  organizers  of  the  company 
obtain  from  the  State  a  charter.  They  then  have  a  formal  meet- 
ing, at  which  they  adopt  a  set  of  laws  for  the  guidance  of  the 
company,  and  elect  directors,  who  are  to  manage  the  company. 
The  directors  then  proceed  to  the  election,  from  their  number,  of 


384  THE    METALLURGY 

corporate  officers  of  the  company.  These  officers  are  the  president, 
vice-president,  secretary  and  treasurer.  The  directors  also  ap- 
point a  manager,  who  has  direct  active  charge  of  the  affairs  of 
the  company.  Either  he,  or  the  directors,  then  appoint  a  super- 
intendent who  attends  to  the  technical  affairs  of  the  plant,  while 
the  manager  gives  his  chief  attention  to  commercial  or  business 
matters. 

The  organization  of  the  technical  force,  at  the  head  of  which 
is  the  superintendent  (sometimes  chief  engineer),  may  include; 
(i)  the  supply  or  purchasing  department;  (2)  labor  force;  (3) 
accounting. 

(1)  The  supply  department  attends  to  the  purchase,  care  and 
issuing  of  ores,  fuel,  supplies  of  all  sorts,  fluxes  and  chemicals; 
also  to  the  disposal  of  the  products  of  the  works. 

(2)  The  labor  force  includes  the  general  foreman,  under- fore- 
men, labor  for  repairs  and  improvements,  force  of  the  laboratory 
and  assay-office. 

(3)  The  accounting  department  attends  to  the  accounting,  pay- 
roll, cost-keeping  and  distribution  of  costs. 

Referring  to  the  question  of  costs,  they  may  be  divided  into 
(x)  prime  cost  or  direct  treatment  and  (y)  general  expense  or 
fixed  charges.  The  prime  cost  varies  almost  directly  as  the  ton- 
nage, while  general  expense  is  a  charge  which  remains  much  the 
same  whatever  the  speed  of  operation,  as  for  instance,  office  ex- 
penses, taxes,  insurance,  etc.  Costs  vary  as  follows : 

The  cost  of  fuel,  per  ton  of  ore,  diminishes  as  the  output  in- 
creases. 

The  percentage  of  labor  to  the  total  cost  is  at  a  minimum  in 
years  of  average  activity.  It  increases  in  dull  and  in  prosperous 
years.  At  a  prosperous  time,  to  meet  the  increased  demand,  extra 
men  must  be  hired  who  do  not  work  to  the  same  advantage  as  the 
force  just  sufficient  for  effective  operation. 

Fixed  charges  or  general  expense  includes: 

(a).  General  expense  of  the  works,  salaries  of  technical  men 
in  charge. 

(b).  General  expense  of  operating,  management,  insurance, 
office  force,  customs  duties,  indemnities  and  damages,  hospital 
and  sickness. 


OF    THE    COMMON    METALS. 


385 


(c).  General  expense  of  the  company.  Dividends  and  interest 
on  bonds,  general  administration,  general  accounts  and  sales  ex- 
penses. (This  item  (c)  is  often  estimated  at  10%  of  the  work- 
ing costs). 

In  connection  with  this  matter  of  costs,  there  are  yet  two  con- 
siderations, which  have  a  bearing  on  the  returns  on  capital  in- 
vested. These  are  depreciation  of  plant,  and  interest  on  invested 
capital.  A  plant  will  depreciate  in  value  at  the  rate  of  10  to  15% 
annually,  so  that  even  if  kept  in  repair,  by  the  end  of  10  years  it 


FIG.  171.     ELECTRIC  TROLLEY  SYSTEM. 


may  be  regarded  as  having  no  value.  If,  however,  a  company 
sets  aside  from  its  earnings  an  amount  sufficient  to  put  in  needed 
improvements  and  so  keep  the  plant  up  to  date,  then  depreciation 
may  be  neglected.  In  case  this  is  not  done  then  dividends  must 
be  sufficient,  not  only  to  pay  interest  on  the  capital  invested,  but 
also  to  meet  depreciation.  Otherwise  the  investor  will  never  re- 
cover his  principal. 

There  are  two  kinds  of  metallurgical  works,  mines  and  custom. 


386  THE    METALLURGY 

A  mines  works  gets  its  supplies  from  a  mine  of  the  same  com- 
pany and  treats  the  ore  which  is  delivered  to  it.  In  such  a  case 
it  does  not  have  to  pay  outright  for  the  value  of  the  ore,  but 
performs  the  duty  of  reducing  or  winning  it  ready  for  market. 

On  the  other  hand,  a  custom  works  finds  the  buying  of  ores  the 
most  serious  item  of  expense,  especially  in  the  case  of  precious- 
metal  ores.  Its  necessary  supply  depends  upon  its  distance  from 
the  mines  and  also  upon  the  certainty  of  supply.  In  the  case 
of  the  Lake  iron-ore  supply,  for  example,  a  stock  must  be  accum- 
ulated by  the  beginning  of  winter  to  keep  the  furnaces  going 
until  the  opening  of  navigation,  that  is  for  a  period  of  six  months. 
The  smelting  wrorks  in  the  Western  Rocky  Mountain  States  carry 
a  supply  of  from  two  to  six  weeks.  In  milling,  where  a  mine  is 
close  by,  the  supply  may  be  replenished  as  fast  as  used,  so  that 
only  enough  for  one  or  two  days'  running  is  needed. 

Besides  ore,  fuel  is  an  important  consideration,  both  for  smelt- 
ing and  for  power!  Where  it  is  hard  to  get,  water  is  one  of  the 
most  important  items  in  milling.  Supplies,  such  as  mercury, 
chemicals,  oil,  grease,  waste,  tools,  castings  and  miscellaneous 
supplies,  cut  more  or  less  of  a  figure,  as  for  example,  in  cyaniding 
ore,  the  potassium  cyanide  may  be  the  largest  single  item. 

1 20.     INVESTMENT  REQUIRED  ON   ORIGINAL   PLANT. 

1.  As  a  preliminary  to  undertaking  metallurgical  operations,  an 
investigation  must  be  made,  not  only  of  the  process  and  plant, 
but  of  all  the  limiting  conditions.    This  includes  supplies,  markets, 
railroad  facilities  and  statistics,  labor  and  reliable  civil  conditons, 
and  an  estimate  of  profits  which  can  be  expected. 

2.  Next  comes  the  promoting  of  the  enterprise,  or  interesting 
capital  to  erect  and  operate  the  required  plant.     This  work  goes 
on  concurrently  with  the  organization  of  the  operating  company. 

3.  Often  the  promoters  acquire  the  necessary  real  estate  and 
the  rights  which  go  with  it.     Provision  is  made  for  access  by 
the  railroads,  and  the  necessary  trackage,  and  for  common  roads 
Not  only  must  water  and   power  be   obtained,   but  also   rights 
of  way  for  them  to  the  property.     Quarry  rights  are  to  be  ac- 
quired, and  a  proper  supply  of  fluxes  provided. 


OF    THE    COMMON    METALS.  387 

4.  Then  come  hydraulic  works  if  needed,  and  the  building  of 
dam,  canal,  reservoir,  water  power,  transmission  line,  and  the 
putting  in  of  telegraph,  telephone  and  mail  facilities.    For  a  long 
transmission  line  there  are  substations  and  the  distributing  sys- 
tem, including  engineers  for  designing*  and  supervision. 

5.  There  must  not  be  forgotten  the  salaries  of  officers  of  the 
company  who  are  to  receive  pay  during  construction,  nor  that  all 
money  must  be  accounted  for  and  the  costs  reckoned  up  by  a 
skilled  accountant      We  may  also  include  the  interest  on  money 
expended  in  the  preliminary  construction  stage,  or  until  the  plant 
is   completed   and   earning   money  to   defray   running   expenses. 
When  in  full  operation,  any  further  money  needed  should  come 
out  of  the  gross  profits. 

Construction. — In  beginning  this  the  costs  should  be  worked  out 
in  great  detail,  good  materials  must  be  obtained,  and,  as  above 
stated,  working  capital  must  be  enough  to  keep  things  going  until 
returns  come  in.  In  the  design  of  the  plant  there  should  be  dupli- 
cation of  parts,  so  that,  in  case  of  break-downs,  there  is  no  in- 
terruption of  operations. 

Operating  expense. — This  may  be  divided  into  two : 

Fixed  charges  or  general  expense,  including  insurance,  taxes, 
improvement  fund,  office  and  miscellaneous  charges. 

Prime  costs,  into  which  comes  labor,  supplies  and  repairs. 
Fixed  charges  remain  much  the  same,  whether  the  plant  is  in  full 
operation  or  not ;  but  prime  cost  varies  according  to  the  tonnage 
treated. 

On  the  business  side  of  the  enterprise  come,  often  the  disposal 
or  sale  of  the  products,  the  accounting  and  estimating  of  costs. 

121.     PROFITS. 

Profits  arising  from  the  operation  of  a  plant,  whether  inde- 
pendent (custom),  or  connected  to  a  mine,  may  be  defined  as  the 
difference  between  the  total  gross  costs  and  the  returns  on  the 
metal  product  sales.  These  costs  may  be  improved,  either  by 
better  extraction  or  recovery  from  the  ore,  or  by  lower  costs  of 
treatment,  due  to  methods  involving  a  saving  in  labor,  supplies 
or  fuel. 


388  THE    METALLURGY 

The  profits  of  customs  works  are : 

1.  Profits  on  treatment. — A  definite  treatment  charge  is  made 
by  the  works  for  the  ore.    The  difference  between  this  charge  and 
the  actual  gross  cost  represents  the  treatment  profit.    In  the  early 
days  of  smelting  in  Leadville  charges  of  $60  per  ton  were  made 
when  the  actual  cost  was  $20.     The  difference  represented  the 
profits  per  ton.    Nowadays  $i  to  $2  per  ton,  because  of  competi- 
tion, must  satisfy  such  custom  works. 

2.  Profit  on  extraction. — A  charge  is  made  against  the  ore  by 
which  a  guaranteed  extraction  is  made,  which  is  deducted  from 
the  gross  values  in  the  contained  metals.    This  is  supposed  to  meet 
the  loss  incurred  in  treatment,  but  as  a  matter  of  fact,  the  ex- 
traction may  be  better  than  this,  and  thus  a  profit  is  made  by  the 
reduction  works.    For  example  we  have  on  a  number  of  different 
lots  of  ores  the  following  estimate : 

Average  treatment  charges  per  ton ....   $6.25 
Actual  cost  of  treatment 4.50 


Leaving  a  profit  on  treatment  of $:-75 

Gain  in  extraction  over  the  prescribed  amount: 

Gain  in  silver  per  ton $0.25 

gold          "       o.io 

lead  "      0.35 

0.70 


Total  gain  or  profit $2.45 

In  milling,  this  calculation  remains  true  whatever  the  charac- 
ter of  the  ore ;  in  smelting,  however,  a  silicious  ore,  which  would 
need  much  flux,  would  yield  less  profit,  because  fluxes  would  be 
taking  the  place  of  profit-bearing  ore.  The  slag  of  such  ore, 
being  considerable  in  amount,  carries  off  values  according  to  its 
weight. 

The  following  figures  represent  profits  yielded  by  an  ore  of  a 
company  owning  its  own  mine,  the  Robinson  Co.,  on  the  Rand, 
South  Africa. 


OF    THE    COMMON    METALS.  389 

Gold  recovered  at  the  stamps $20.50 

"  "  by  cyanide    5.70 


$26.20 

Cost  of  mining $6.55 

"     of  milling   0.98 

"     of  cyaniding    0.97 

Net  profits  per  ton J7-9O 


$26.40 

Besides  the  costs  of  operation,  as  above  given,  we  have  to 
consider  the  following  costs,  which  are  often  not  considered  in 
estimates : 

Expressage  on  gold  and  silver  bars  to  the  mint  or  to  market. 

Costs  of  selling,  generally  a  percentage  on  the  gross  amount 
of  the  sale. 

Interest  on  bonds,  where  the  company  has  issued  them.  This 
comes  out  of  the  profits  before  stockholders  can  expect  any  divi- 
dends. 

Sinking-fund  reserved  to  repay  bonds.  This  is  an  annual  de- 
duction, which,  with  interest  obtained  by  investing  the  accruing 
amounts,  will  be  enough  at  maturity  to  meet  the  obligation. 

General  office  expense.  This  has  reference  to  the  home  office, 
where  there  is  one,  generally  in  some  large  city.  The  object  of 
such  an  office  is  that  the  stockholders  may  be  in  touch  with  the 
property  in  which  they  have  invested. 

Salaries  of  officials,  such  as  corporation  officers. 

Insurance  upon  the  plant,  and  on  the  metals  in  transit. 

Taxes,  including  both  general  and  local  taxes. 

Royalty,  where  an  amount  must  be  paid  out  for  the  use  of  a 
patented  metallurgical  process. 

122.     ORGANIZATION. 

Mines  in  a  given  country  have  value  only  when  the  ore  has 
been  extracted  and  reduced.  This  means  the  employment  of 
labor.  For  example :  A  new  camp  has  been  prospected,  and 


39°  THE    METALLURGY 

upon  promise  of  a  future,  works  are  erected  to  treat  the  ores  of 
that  camp.  A  century  or  more  ago  the  methods  of  reduction  were 
very  simple,  needing  but  little  capital,  and  labor,  and  with  a  small 
output,  requiring  only  the  simpler  methods  of  transportation. 
Upon  the  skill  of  the  workmen  depended  the  success  of  the  oper- 
ation. 

Now,  however,  this  is  profoundly  changed.  Small  furnaces 
are  replaced  by  large  ones,  and  in  place  of  operations  dependent 
on  the  skill  of  the  workmen,  the  metallurgist  computes  his  charge. 
To  manual  skill  has  been  added  intellectual  skill,  in  order  to  in- 
sure the  certain  and  exact  operations  of  the  furnace.  Where,  at 
first  by  hard  labor  the  puddler  worked  up  his  charge  of  iron, 
today  he  operates  levers  while  watching  operations.  In  place  of 
muscular  effort  he  must  apply  intelligent  direction. 

The  superintendent,  the  assistant  or  the  foreman  must  not  only 
be  informed  as  to  the  actual  technical  operations,  but  he  must 
also  know  how  to  arrange  and  to  organize.  And  in  addition,  he 
must  be  able  to  handle  men  effectively.  He  must  possess  tact,  dis- 
cretion and  firmness.  He  must  be  strict  but  just,  able  to  encour- 
age as  well  to  drive. 

The  choice  of  men,  who  are  to  run  a  plant,  is  often  the  im- 
portant point  in  the  success  of  an  enterprise.  In  certain  cases  it 
has  been  the  custom  to  put  in  men  who  have  the  "pull-'  because 
their  relatives  or  friends  are  on  the  directorate.  Thus  success  is 
dampened  at  the  start.  We  train  men  at  mining  schools  in  prep- 
aration for  technical  duties.  Not  that  they  can  assume  responsi- 
bility without  practice,  but  that,  knowing  principles  and  having 
the  mental  training,  they  are  fitted  to  acquire,  to  advance,  and 
finally  to  be  given  responsibility. 

The  pay  of  intellectual  work,  as  for  the  office  force,  superin- 
tendent, assayer,  chemist,  and  foremen,  is  monthly.  In  certain 
cases  an  annual  premium  has  been  paid.  However,  in  those  plants 
where  the  exhaustion  of  a  mine  or  a  camp  may  cause  the  closing 
of  a  work,  such  premiums  can  hardly  be  assured.  These  methods, 
however,  often  give  good  results,  especially  where  this  premium 
depends  on  tonnage,  or,  in  the  case  of  the  superintendent,  on  the 
lessening  of  costs.  There  is  a  big  difference  between  the  man  who 
will  draw  his  salary  anyhow,  and  one  who  knows  he  is  to  get  more 


OF    THE    COMMON    METALS.  391 

when  he  exerts  himself  to  get  to  work  promptly  in  the  morning, 
or  who  stops  some  leaks  which  otherwise  he  would  let  go. 

The  quality,  as  well  as  the  quantity  of  work,  has  to  be  seriously 
considered.  In  zinc  plants,  for  example,  success  is  largely  depend- 
ent on  a  high  extraction,  due  to  skill  in  retorting.  This  extraction 
may  easily  drop  from  3  to  5%  through  carelessness,  or  even  by 
lack  of  determination  to  get  the  best  possible  quantities  out  of  the 
ore.  In  roasting  and  in  puddling,  skill  in  firing  means  a  saving 
of  fuel  out  of  all  proportions  to  the  pay  of  the  fireman,  a  fact 
which  we  better  know  in  connection  with  firing  a  boiler.  Again, 
about  a  silver-lead  works,  good  feeding  and  good  furnacing  counts, 
and,  more  than  all,  skillfully  compounded  charges  of  uniform  com- 
position. There  is  this  much  about  the  charging  of  such  a  fur- 
nace, that,  if  it  is  correct,  it  will  not  only  give  the  best  reduction, 
but  also  the  smoothest  running. 

Let  us  here  consider  the  matter  of  cleanliness  and  neatness 
about  the  plants,  especially  where,  as  in  refineries,  high-grade  ma- 
terials are  treated.  Since  these  materials  are  treated  on  a  close 
margin  of  profit,  and  since  a  small  percentage  loss  will  turn  a 
profit  into  a  loss,  it  is  all-important  to  look  closely  into  the  matter 
of  preventing  small  losses.  Valuable  materials  such  as  drosses, 
skimmings,  spillings  from  retorts,  and  from  cupel  furnaces,  flue- 
dust  and  sweepings,  should  be  carefully  and  anxiously  gathered  up 
and  preserved  to  go  back  for  recovery  or  re-treatment.  One  may 
find  in  accumulations  of  ashes  valuable  concentrate,  containing 
several  hundred  dollars  per  ton,  these  ashes  being  accumulated 
from  the  retorts  of  a  refinery.  About  mints  and  jewelers  estab- 
lishments several  hundred  dollars  at  a  time  have  been  obtained 
by  sweeps  and  from  the  ashes  of  old  carpets,  which,  after  having 
been  for  some  years  upon  the  floor,  have  been  burned. 

In  general,  the  modern  practice  has  been  to  give  more  and  more 
attention  to  dust-losses,  especially  in  connection  with  blast-fur- 
naces which  produce  strong  air  currents.  This  means,  however, 
large  investments  in  flues  and  stacks.  It  is  well  to  remember, 
that  for  low-grade  materials  the  same  rule  would  not  apply, 
since  sometimes  the  flue-dust  recovered  from  a  low-grade  ore  will 
hardly  pay  the  cost  of  re-treatment  plus  the  interest  on  the  flue- 
chamber  needed. 


392  THE    METALLURGY 

In  certain  respects  the  labor  about  a  metallurgical  works  dif- 
fers from  that  about  other  industries.  The  work  is  carried  on  by 
a  crew  who  work  together,  each  man  having  a  part  of  the  opera- 
tion dependent  on  him  alone,  all  being  directed  by  a  foreman,  who 
sees  to  the  regular  operation  of  the  apparatus.  Thus,  about  a 
silver-lead  blast-furnace  we  have  the  feeder  and  assistant,  the 
weigher,  the  wheelers  or  trammers  who  bring  in  the  fuel  and 
charges — all  men  who  work  on  the  feed-floor.  Below  there  is  the 
furnace-man,  the  tapper,  the  pot-pushers,  each  having  a  particular 
duty.  These  are  called  inside  men.  They  work  in  shifts  of  12 
hours,  and  one  crew  replaces  the  other,  while,  if  there  is  a  little 
delay  in  the  arrival  of  any  man,  his  partner  must  await  his  com- 
ing or  until  he  is  relieved.  These  men  are  by  no  means  paid  the 
same;  for  example,  the  furnace-man  or  the  feeder  is  paid  more 
than  the  other  men,  and  the  wage  varies  according  to  the  skill 
and  knowledge  required  of  the  men. 


122.     GENERAL  REMARKS  ox  MANAGEMENT  AND  LABOR. 

Engineering  is  the  art  of  doing  with  one  dollar  what  any 
bungler  after  a  fashion  can  do  with  two.  In  other  words,  a  com- 
petent man  is  necessary  to  the  best  results,  and  with  a  small  mar- 
gin of  profits  to  count  on,  he  is  absolutely  essential. 

The  capacity  and  efficiency  of  a  plant  depend  on  the  intelligence 
and  reliability  of  men  in  its  different  departments.  You  must 
choose  your  men  with  this  in  view.  That  is,  the  lower  grade 
labor  is  to  be  used  where  routine  and  hard  work  come  in ;  the 
higher,  where  judgment  is  needed.  While  lower  grade  labor  is 
faithful  it  is  stupid  and  may  make  blunders.  However,  if  a  man 
has  been  supervised  and  trained,  then  he  can  be  trusted  if  his 
nature  is  trustworthy.  So  cheap  labor,  as  in  supervising  ma- 
chines, may  be  offset  by  losses  or  actual  disaster.  In  employing 
men  it  is  best  to  mix  the  nationalities  to  avoid  the  clannishness 
which  may  result  in  the  men  of  the  same  nation  combining  against 
their  employers  or  bosses. 

Though  the  men  are  generally  left  to  shift  for  themselves,  cer- 
tain sanitary  precautions  regarding  them  are  well  repaid. 


OF    THE    COMMON    METALS.  393 

The  manager  or  superintendent  at  a  works  should  give  orders 
through  the  foreman,  otherwise  discipline  is  injured,  and  the  fore- 
man's well-laid  plans  may  be  disarranged.  If  prompt  change  is 
imperative  you  may  make  it  but  you  should  at  once  hunt  up  the 
foreman  and  tell  him  of  it. 

Beginning  with  the  manager,  who  takes  from  the  company  the 
responsibility  of  operations,  we  find  that  he  must  be  a  man  espe- 
cially informed  in  a  business  way,  and  in  the  business  of  that 
branch  of  metallurgy  of  which  he  has  charge.  He  appoints  a 
superintendent,  who  has  direct  charge  of  the  technical  operations, 
while  the  manager  gives  himself  to  the  care  of  the  business 
matters. 

The  superintendent,  often  also  the  metallurgist,  has  direct  con- 
tact with  the  furnaces  and  metallurgical  machinery.  When  mat- 
ters go  wrong  it  is  he  who  is  called  on  to  make  them  right  at 
no  matter  what  hour  of  the  day  or  night.  Is  a  furnace  in  bad 
condition,  or  has  a  machine  broken  down,  he  is  responsible.  When 
all  is  going  smoothly  his  duties  may  be  light,  but,  when  troubles 
occur  and  the  works  is  losing  money,  then  is  the  time  when  he 
must  work  hard.  If  he  fails  in  adjusting  matters  satisfactorily, 
no  excuses  can  be  accepted ;  he  must  succeed  or  resign.  Much  of 
his  success  depends  upon  his  subordinates,  among  whom  we  must 
first  place  the  foreman. 

While  the  foreman  is  not  expected  in  lead-silver  smelting  opera- 
tions to  make  changes  in  the  charge,  or,  in  any  case,  to  devise  new 
schemes  of  operation,  his  suggestions  are  apt  to  be  valuable,  and 
upon  the  basis  laid  down,  he  has  considerable  liberty  of  action. 
His  skill  comes  in  particularly  in  the  handling  of  the  men,  whom 
'he  must  not  only  distribute  properly  to  do  the  work,  but  must 
also  see  that  they  keep  busy  and  working  efficiently. 

Qualities  of  a  good  foreman.  In  quizzing  an  applicant  for  such 
a  position  make  a  note  of  his  former  employers,  ask  him  if  he 
gets  on  well  with  his  men,  if  he  scolds  them,  if  he  has  been  threat- 
ened by  them,  or  on  the  contrary  if  he  treats  them  "white,"  or 
if  he  shows  them  how  to  work  by  doing  it  himself  instead  of  work- 
ing all  of  them.  A  foreman  must  not  be  thought  a  good  fellow 
by  his  men.  On  some  jobs  a  good  foreman  must  watch  and  fore- 
see everything  to  be  sure  that  nothing  is  neglected,  and,  in  such 


394  THE    METALLURGY 

a  case,  he  has  less  need  to  be  a  strict  disciplinarian.  On  other 
work,  and  in  order  to  drive,  especially  when  he  has  many  men  to 
handle,  he  must  be  stern  but  just.  He  cannot  hobnob  or  be 
familiar  with  the  men,  but  must  have  reserve  of  manner  though 
looking  after  their  welfare.  Men  work,  not  to  please  the  foreman, 
but  because  they  need  the  money ;  in  general,  money  talks. 

The  chemist  and  assayer  is  called  on  principally  to  give  results 
with  promptness  and  accuracy.  Besides  his  routine  work  his 
skill  can  be  shown  to  advantage  in  investigation  under  the  direc- 
tion and  advice  of  the  metallurgist. 

The  office  force  has  to  attend  to  the  purchase  and  issuing  of  sup- 
plies, to  the  accounting,  and  to  the  questions  of  costs  and  profits. 

The  various  mechanics  —  as  blacksmith,  carpenter,  engineer, 
machinist — have  not  only  the  repairs,  but  also  the  improvements 
of  the  plant  to  attend  to,  under  the  personal  direction  of  the  super- 
intendent, who  may,  where  the  work  requires  it,  employ  a  drafts- 
man and  engineer  to  attend  to  construction.  It  is  a  rule  that  when 
in  an  emergency  the  foreman  calls  for  work  to  be  done  in  a  hurry, 
then  such  work  has  precedence. 

Skilled  and  unskilled  labor.  The  lo-hour  labor  about  reduction 
works  can  be  performed  largely  by  unskilled  men,  called  outside 
laborers  or  roustabouts.  They  perform  work  involving  the  use  of 
pick  and  shovel,  such  as  unloading  cars,  handling  the  products 
of  the  works  and  assisting  in  construction  and  repairs. 

The  skilled  labor,  in  shifts  of  8  to  12  hours,  called  also  inside 
men,  receive  extra  pay  according  to  the  skill  needed,  and  the 
longer  hours.  They  differ,  also,  in  the  fact  that  they  are  responsi- 
ble for  the  successful  performance  of  the  duties  given  them,  and 
they  are  expected  to  work  until  they  are  relieved  by  their  partners, 
or  until  the  foreman  has  provided  someone  to  take  their  place. 

For  keeping  up  discipline,  and  to  prevent  slackness  in  the  work, 
certain  rules,  the  result  of  experience,  have  been  laid  down  for  the 
guidance  of  the  men.  These  are  : 

Rules  of  shifts. — The  men  must  be  promptly  at  work,  and  work 
full  time.  Ten  hours  runs  from  7  A.  M.  to  12  M.  and  from  I  P.  M. 
to  6  P.  M.  in  summer,  while  in  winter,  the  noon  hour  is  shortened 
to  thirty  minutes. 


OF    THE    COMMON    METALS.  395 

Inside  or  twelve-hour  men  must  be  on  hand  all  the  time  and 
eat  their  luncheon  when  they  get  the  chance.  They  must  always 
be  on  watch  where  they  are  needed.  Charge-wheelers  must  keep 
up  the  supply,  but  otherwise  they  may  rest  at  intervals,  and  are  not 
to  be  called  on  to  do  other  work  than  to  sweep  up  before  going  off 
shift.  The  inside  man  may  go  off  when  relieved  by  his  partner 
and  must  wait  for  his  partner  to  come.  If  the  latter  does  not 
come  the  foreman  provides  another  man,  who  then  takes  first 
place,  the  absent  man  losing  his  place  unless  he  has  a  good  excuse 
for  his  absence.  When  such  a  man  is  to  be  unavoidably  detained, 
or  is  sick,  he  is  to  notify  the  foreman  by  message,  who  will  supply 
his  place.  When  the  man  wishes  to  return  to  work  he  must 
notify  the  foreman  one  shift  in  advance,  so  that  his  substitute  is 
not  put  out  of  a  job  for  which  he  had  prepared. 

When  men  get  sick  on  shift,  try  to  hold  them  to  the  end  of 
the  shift  under  the  idea  that  it  ij  hard  to  fill  their  places  at  such 
short  notice. 

Labor  troubles.  Men  must  obey  orders,  and,  if  not,  must  be 
discharged,  irrespective  of  whom  they  may  be. 

Be  strict  and  just.  Some  employers  find  it  well  to  occasionally 
let  out  a  man.  If  you  run  on  for  a  while  without  doing  so,  begin 
to  be  anxious  for  yourself.  Do  not  trust  a  man  to  do  work  with- 
out supervision,  since  he  may  do  it  wrong,  or  may  get  careless. 

Provision  for  the  care  of  the  men  in  case  of  sickness,  such  as 
a  hospital  or  medical  attendance,  is  often  made.  The  men  pay  for 
hospital  dues  one  dollar  per  month,  and  are  entitled  to  the  privil- 
eges thus  paid  for.  When  a  man  has  worked  to  exceed  five  days 
in  any  given  month  this  one  dollar  is  taken  from  his  pay  for  hos- 
pital fee.  While  accidents  will  occur,  yet  particular  care  must 
be  given  by  the  foreman  to  the  safety  of  the  men,  who  are  apt 
to  get  careless,  and  must  be  warned.  Any  carelessness  on  the 
part  of  the  foreman  in  this  regard  renders  the  company  liable  for 
damages. 
American  as  compared  ivith  Mexican  or  similar  labor. 

American  labor  is  characterized  by  great  intelligence,  energy 
and  responsibility  as  compared  with  Mexican.  Mexicans  who 
have  been  trained  to  such  duties  often  make  excellent  inside  men. 


THE    METALLURGY 

When  well  bossed  or  directed,  Mexicans  do  pretty  well,  though 
often,  from  being  insufficiently  fed,  their  physical  endurance  is 
low.  They  need  watching,  also,  to  prevent  their  idling.  They 
must  be  directed,  as  they  are  apt  to  work  to  disadvantage  on 
account  of  their  thoughtlessness.  It  is  well  to  treat  them  as 
grown  children,  and  to  have  patience  with  them,  but,  at  the  same 
time,  judicious  strictness  is  required. 

In  fixing  upon  a  rate  of  wages,  especially  in  a  new  camp,  it  is 
well  to  have  the  pay  low  to  begin  with.  It  is  always  easy  to  raise 
the  wages  of  the  men,  but  hard  to  reduce  them  once  they  have 
been  fixed.  Among  Mexicans,  whose  tastes  are  simple  and 
whose  wants  are  easily  supplied,  the  tendency  of  high  wages  is 
to  give  more  laying-off  time,  to  the  great  inconvenience  of  the 
works.  This  is  also  the  trouble  of  monthly  payments  by  which, 
when  wages  have  been  paid,  the  men  are  disposed  to  take  a  rest 
to  the  great  detriment  of  steady  running.  To  overcome  these 
troubles  two  methods  have  been  tried.  One  is  of  daily  payments 
by  which  a  Mexican,  who  would  naturally  spend  all  his  money  as 
fast  as  he  gets  it,  has  thus  but  little  to  spend,  and  who  would 
come  round  the  next  day  to  earn  more  ready  cash.  The  other 
system  is  to  pay  him  a  wage,  to  which  is  addecj,  a  premium  which 
increases  with  the  time  worked,  and  is  paid  at  the  end  of  the 
month  if  he  works  the  month  through,  otherwise  not.  This  re- 
ward is  a  bait  which  holds  a  man  to  steady  work. 

If  an  employer  of  Mexican  labor  finds  a  man,  who  excels  in  his 
work,  let  him  avoid  the  danger  arising  from  advancing  his  wages, 
which  often  has  the  effect  of  spoiling  him.  In  speaking  to  and 
directing  a  Mexican  laborer  the  familiar  style,  not  the  formal  or 
polite  one  of  the  books,  is  to  be  used.  The  formal  address  would 
be  a  mistake. 

The  yellow  streak. 

Mexican  and,  in  a  lesser  degree,  other  labor  is  liable  to  show 
its  weakness  and  lack  of  common  honesty  by  pilfering.  So  much  so 
is  this  the  case  in  Mexico  that  a  tool,  or  other  portable  object,  can 
hardly  be  laid  down  but  that  it  will  be  stolen.  Men  must  be  held 
responsible  for  their  tools,  and  such  property  kept  under  lock  and 


OF    THE    COMMON    METALS.  397 

key.  It  is  also  well  to  brand  tools.  In  the  spring,  when  the  pros- 
pector is  about  to  get  out  to  the  hills,  he  finds  it  convenient  to 
draw  on  the  company  in  this  way  for  tools  without  expense  to 
himself. 

In  dealing  with  Mexicans,  however,  the  American  metallurgist 
or  engineer  cannot  afford  to  practise  sharp  tricks.  Whatever 
they  are  themselves,  they  expect  integrity  of  him.  Any  other  pro- 
cedure would  soon  be  known  to  his  vital  disadvantage.  Of 
course,  some  Americans  have  inherited  the  yellow  streak  from 
their  ancestors,  and  their  training  may  have  also  contributed  to 
the  same  result.  Even  they  should  be  wise  enough  to  see  that 
they  must  put  restraint  on  themselves,  and  realize  the  degradation 
arising  from  such  meanness.  The  mining  engineer  is  trusted  with 
the  control  of  enterprises,  and  with  the  care  of  the  precious  metals 
which  he  produces.  His  success  depends,  not  only  on  good  reso- 
lutions, but  also  in  keeping  them. 

Mines  have  a  practical  value,  only  when  they  have  been  devel- 
oped so  that  it  is  possible  to  know  what  the  future  has  to  offer 
in  erecting  reduction  works.  Where  the  value  of  the  ore  justifies 
it,  the  beginnings  of  metallurgical  operations  should  be  simple, 
depending  upon  the  skill  of  the  workman  and  not  upon  a  com- 
plicated plant.  As,  however,  knowledge  has  been  obtained,  the 
erection  of  an  efficient  plant  should  be  undertaken. 

It  is  not  sufficient  that  the  superintendent  shall  simply  know 
how  technical  work  should  be  done;  he  must  also  be  able  to 
organize  his  force  and  to  handle  it  efficiently.  To  do  this  he  must 
have  tact  as  well  as  knowledge.  The  right  men  make  the  success 
of  the  enterprise.  For  all  this,  influence  or  pull  often  desides  who 
is  to  control,  and  may  result  in  the  choice  of  an  incompetent,  un- 
trained, or  inexperienced  man,  thus  inviting  disaster. 

The  tendency  at  modern  mills  and  metallurgical  works  is  to 
reduce  the  labor  per  ton  needed  by  means  of  mechanical  con- 
veniences, so  that  the  labor  of  the  individual  becomes  less  strenu- 
ous, less  trying  on  his  strength  and  endurance,  and  yet  he  is  better 
paid.  As  a  result  he  is  better  able  and  more  anxious  to  retain  his 
job,  and  his  services  are  more  to  be  relied  on  than  is  the  case 
with  cheaper  labor. 


398  THE    METALLURGY 

1.  With  respect  to  the  labor  needed,  a  large  mill  is  run  with 
proportionately  less  labor  than  a  small  one. 

2.  An  easily  treated  ore  needs  proportionately  less  labor,  since 
there  are  fewer  operations  and  the  care  needed  is  less. 

3.  Eight-hour  shifts  need  one-half  more  men  than   1 2-hour 
shifts. 

4.  Again,  steam  power  needs  more  labor  than  water  power 
where  often  it  is  but  the  question  of  an  occasional  adjustment  of 
a  valve. 

5.  Labor-saving  machinery,  provided  it  is  reliable,  makes  a 
difference  in  costs,  but  it  must  not  be  forgotten  that  the  saving  of 
labor  must  not  be  at  the  expense  of  efficiency,  and,  also  the  ques- 
tion of  "how  much"  often  comes  in,  so  that  in  certain  cases  it  is 
not  worth  while  to  put  in  the  labor-saving  appliance. 

6.  Cost  and  quality  of  labor.     Cheap  labor  is  needed  in  larger 
quantity,  and   Mexican  labor  needs  much   supervision  by  gang 
bosses.    Cheapness  of  living  also  effects  labor. 

Men  at  an  amalgamating  silver  mill  (100  stamps). 

Foreman  Head  amalgamator 

Millwright  Four  amalgamators 

Pipe  fitter  Two  crusher  men 

Two  engineers  Two  oilers 

Two  firemen  Two  feeders 

Night  foreman  Two  laborers 

In  starting  a  new  metallurgical  works  a  list  of  men  and  their 
places  or  occupations  should  be  arranged,  and  as  men  are  en- 
gaged they  are  assigned  on  the  list.  On  the  day  of  starting, 
others  apply  who  must  be  assigned  rather  hastily.  All  these  men 
must  be  questioned  as  to  their  qualifications,  and  given  places 
they  can  fill.  About  the  time  a  new  works  is  to  be  started,  men 
skilled  in  the  operation  of  the  plant  are  often  willing  to  accept 
common  labor  while  waiting  for  the  start. 

The  men  needed  and  the  cost  of  labor  for  a  single  blast-furnace 
lead-smelting  plant  per  shift  of  12  hours  is: 


OF    THE    COMMON    METALS.  399 


Inside  men  : 

i   feeder   .........................  $2  .  50 

i   feeder's  helper    .................    i  .  80 

3  charge-  wheelers  @  $1.80  .........   5  .40 

i    weigher    .......................    2.4.0 

i    furnace-man    ...................    2  .  50 

1  tapper   .........................    i  .  80 

2  pot  pushers  @  $i  .  80  ............    3  .  60 

i  engineer   .......................   3  .  50 

i   foreman    .......................   4  .  25 


This  is  based  upon  common  labor  being  $1.50  per  day  of  10 
hours.     When  labor  is  more  expensive  these  prices  increase  pro- 
portionately. 
Ten-hour  men  : 

i  dump  man   .....................  $i  .  50 

4  sampling  mill  men  @  $1.50  ......   6.00 

i   sampling   foreman    ..............   2  .  50 

i  carpenter    ......................   3  .  oo 

i  blacksmith    .....................    3  .  50 

$16.50 


Total  cost  for  labor $72.00 

This  is  a  minimum  of  labor,  since  improvements  and  emergency 
work  calls  for  other  labor.  It  is  well  to  have  such  extra  men, 
who  can  be  called  on  to  supply  places  caused  by  sickness,  etc. 

In  gold  stamp-milling  the  duties  of  the  mill  men  in  a  large 
mill  are: 

Foreman. — He  has  general  supervision  of  the  mill  and  looks 
after  the  handling  and  cleaning  of  all  the  amalgan  collected. 

Amalgamators. — They  dress  the  chuck-blocks  and  plates  and 
keep  them  in  good  condition.  They  set  tappets,  regulate  water- 
supply  and  make  all  renewals. 

Feeders. — Attend  to  the  uniform  feeding  of  the  batteries  and 
assist  the  amalgamators  in  renewals  and  on  the  clean-up.  A 
good  feeder  is  a  valuable  man  about  the  mill. 


4OO  THE    METALLURGY 

Vanner-men. — They  attend  to  the  vanners  or  tables.  They 
must  be  men  with  a  good  deal  of  experience.  In  fact  they  must 
first  serve  as  helpers  about  the  tables  as  sulphide-pullers. 

Crusher-men  operate  the  crushers  for  the  coarse  ore  entering 
the  mill. 

Oilers  oil  all  the  machinery. 

Sulphide-pullers  remove  the  concentrate  or  sulphides  from 
the  vanner  boxes,  and  store  them  ready  for  shipment. 

Engineers  run  the  power  plant  and  have  charge  of 

Firemen,  who  fire  the  boilers  and  remove  ashes. 

Coal-passers,  who  bring  in  coal  to  the  boilers. 

On  repairs  there  are  carpenters,  and  laborers  who  assist  them 
on  the  vanners.  This  repair  force  gets  the  help  of  a  special  van- 
ner-man. 

At  the  Treadwell  240  stamp-mill  at  Douglas  Island,  Alaska, 
the  following  men  are  needed  each  24  hours. 

1  foreman  (12  hours),  $150  per  month $  5.00 

4  amalgamators  (12  hours),  $90  per  month.  ...    12.00 

8  feeders   (12  hours),  $70  per  month 18.64 

4  vanner-men  (12  hours),  $65  per  month 8.68 

2  oilers   (12  hours),  $65  per  month '    4.34 

2  sulphide-pullers  (10  hours),  $2  per  day 4.00 

2  sulphide-shovelers  do  hours),  $2  per  day.  ...     4.00 

2  engineers  (12  hours),  $2.50  per  day 5.00 

2  firemen  (12  hours),  $2.50  per  day 5.00 

2  coal-passers  (  10  hours),  $2  per  day 4.00 

4  crusher-men   (10  hours),  $2.25  per  day 9.00 


Making  a  daily  cost  for  operating  labor  of $79.66 

Repairs. — 

i   carpenter  (  10  hours),  $4  per  day $4.00 

i  vanner-man  (12  hours),  $100  per  month 3-33 

i  laborer  (10  hours),  $2  per  day 2.00 


Total  labor  cost  for  repairs    $9-33 

To  these  costs  must  be  added  that  for  board  and  lodging  fur- 
nished by  the  company. 


OF    THE    COMMON    METALS.  4OI 

In  a  4O-stamp  silver  amalgamation  mill,  having  24  pans,  and 
with  a  capacity  of  150  tons  daily,  the  labor  in  24  hours  would  be: 

4  pan-men   (12  hours)   @  $4 $16 

2  helpers   ( 12  hours)    @  $3 6 

10  tankmen   ( 12  hours)   @  $3 30 


Making  an  inside  labor-cost  of $52 

These  prices  prevail  in  California  and  in  other  higher-priced 
camps  in  the  Rocky  Mountain  region. 

124.     THE   PURCHASING  OF   ORES   IN   THE   ROCKY   MOUNTAIN 

STATES. 

Smelting  works  purchase  ores  of  every  kind  according  to  their 
requirements,  provided  the  ores  have  value  enough  in  them  to 
pay  for  smelting.  They  are  purchased  according  to  their  con- 
stitution and  upon  a  pre-arranged  schedule,  an  example  of  which 
is  here  given. 

Schedule 

For  Dry  Ores,  Concentrate  and  Tailing,  Lead  Ores  and  Lead 
Concentrate,  Clear  Creek  and  Gilpin  Counties,  Colorado. 

February   i,    1905. 
All  Rates  F.  O.  B.  Cars  Denver. 

Dry  Tailing  and  Concentrate. 
Gold,  $19  per  oz.,  if  0.05  oz.  or  over  per  ton. 
Silver,  95%  of  N.  Y.  quotation  day  of  assay,  if  I  oz.  or  over 

per  ton. 
Copper,  dry  assay  (wet  less  1.5  units). 

5%  or  less   $1.25  per  unit 

Over  5  and  incl.  10%    1.50    " 

Over   10%    1-75    "      " 

10%  silica  basis,  ice  up. 
5%  zinc  basis,  3oc  up. 

When  gross  value  is: 

Not  over  $35  per  ton $3 . 50  treatment 

Over  $35  and  incl.  $80  per  ton 4.00 

Over  $80  per  ton 5 .  oo 

Upon  lots  of  less  than  seven  (7)  tons 5  .00 


4O2  THE    METALLURGY 

Dry  Silicious  and  Copper  Ores. 

Gold,       $19  per  oz.,  if  0.05  or  over  per  ton. 

Silver,     95%  of  N.  Y.  quotations,  day  of  assay. 

Copper,  as  in  schedule  of  cencentrates. 

$8  treatment  charges. 

40%  silica  basis,  5c.  down,  and  ice.  up  to  a  maximum  charge  of 

$i  i  on  ores  not  exceeding  $25  gross  value ;  and  $12.50  on  ore 

exceeding  $25  gross  value. 
5%   zinc  limit,  3oc.  up. 


Oxidized  Irony  Ores. 

Gold,       $19  per  oz.,  if  0.04  oz.  or  over  per  ton. 

Silver,     95%  of  X.  Y.  quotation,  date  of  assay. 

Lead,       2$c  per  unit  for  5%  or  over. 

$2  treatment  charges. 

Neutral  basis,  ice.  a  unit  up. 

Gold,       $19.50  per  oz.,  if  0.05  oz.  or  over  per  ton. 

Silver,     95%  of  N.  Y.  quotation,  date  of  assay. 

Lead,       prices  flat. 

Copper,  $i  per  unit  dry  (1.5%  off  wet)  when  ore  assays  2%  wet. 

Zinc,       limit  10% ,  5oc.  up. 


per  unit,  $8.00  w.  c. 

"       "  7.00  " 

"       "  5.00  " 

4.00 

"       "  4.00  " 

"       "  3.00  " 

"       "  2.50  " 


Xcutral  Scliedulc. 

5  to 

10%   Pb.  incl., 

25C 

Over 

IO     " 

15        " 

25 

" 

15   " 

20          " 

25 

a 

20     " 

25          " 

25 

n 

25     " 

30          " 

30 

a 

30    " 

35 

30 

n 

35    " 

40 

30 

it 

40   " 

45 

32 

11 

45    " 

50       " 

35 

(( 

50 

it 

40 

"  2.00 


"  2.00 


"  2.00 


Neutral  basis,  IDC.  up  or  down. 


OF    THE    COMMON    METALS.  403 

Flat  Schedule. 

5  to  10%   Pb.  inch,                   250.  per  unit,  $12.00  w.  c. 

Over  10    "    15                                        25  10.50  " 

"         15     "     20           "                                           25  8.50  " 

"        20     "     25                                                        25  6.50  " 

"  25  "  30   "           30   "   "  6.00  " 

"  30  "  35               30   "   "  4-50  " 

"  35  "  40   "           30   "   "  3.00  " 

"  40  "  45               32  2.00  " 

"  45  "  50   "            35  2.00  " 

"        50                                                                         40  2.00  " 

Neutral  Schedule  to  be  used  when  it  figures  better  for  the  shipper. 

Lead  Concentrate. 

Gold,       $19  per  oz.  if  0.05  oz.  or  over  per  ton. 

Silver      and  Copper  as  in  lead  ores. 

Lead,       prices  flat. 

Silica,     limit  10%,  loc  up. 

Zinc,       limit  5%,  3oc.  up. 

5  to  10%   Pb.  incl.,                     25c.  per  unit,  $4.75  w.  c. 

Over  10    "    15        "                                 24        "       "  4.00  " 

"         15     "     20           "                                              30           "          "  3.50  " 

"        20     "     25                                                           32  3.25  " 

M     25    "    30                                      35  3-25  " 
Upon   Concentrate  assaying  over   30%   lead,   apply   "Neutral 

Schedule"  or  "Flat  Schedule,"  whichever  figures  the  better  for 

the  shipper.    $19  for  gold. 


Schedule 

For  Ore  and  Concentrate,  Boulder  County  and  Cripple  Creek, 

Colorado. 

February,    1905. 

All  Rates  F.  O.  B.  Cars,  Denver. 
Gold,       $19  per  oz.,  0.05  to  2  oz.  incl.  per  ton. 
19.50  per  oz.,  if  over  2  oz.  per  ton. 


404  THE    METALLURGY 

Silver,     90%  of  N.  Y.  quotations,  if  ore  assays  from  I  to  10  oz. 

per  ton. 
95%  of  N.  Y.  quotation,  if  ore  assays  over  10  oz.  per  ton. 

Up  to  $10  gross  value    $4.00  treatment 

Over       10  to  20  gross  value  5.00 

20    "    30  -  5.50 

30    "    40  6.00 

40    "    50  6.50 

50    "    75          "  7-00 

75    "  100  8.00 

100  9.00 

When  ore  does  not  exceed  $10  gross  value,  3%,  sulphur  limit, 
250.  up  to  a  maximum  charge  of  $2.50  per  ton,  zinc  limit  5%, 
then  3oc.  up. 

When  ore  is  over  $10  gross  values,  no  sulphur  limit,  zinc  limit 
$%,  3oc.  up. 

Lead  Ores. 

Apply  Schedule  for  lead  ores  for  Clear  Creek  and  Gilpin  Coun- 
ties. 

Concentrates,  Lead  or  Dry. 

Apply  corresponding  Schedule  for  Clear  Creek  and  Gilpin 
Counties. 

As  an  example  of  the  use  of  the  above  table  let  us  take  an 
ore  containing  14%  SiO2 ;  6%  Fe ;  11%  Zn ;  4%  Mn;  10%  S; 
21 9^  Pb;  60  oz.  Ag  and  0.2  oz.  Au  per  ton.  The  ore  is  evidently 
a  lead  ore  and  we  will  figure  the  cost  per  ton  on  both  the  neutral 
and  on  the  flat  schedules. 

The  silica  excess  will  be  i4SiO2  less  (6Fe  -f  4Mn)  =  4% 
SiO2  excess. 

The  zinc  being  11%,  its  excess  of  i%   over  the  limit,  makes 
a  penalty  of  5oc  to  be  added  to  the  cost  of  treatment. 
On  the  neutral  schedule 

Gold  0.2  oz.  @  $19.50   $  3.90 

Silver  60  oz.  @  95%   of,  say,  62c  (N.  Y.  quota- 
tion)         35.34 

Lead  21%    for  units)  @  25c 5.25 


Gross  metal  values $44 . 49 


OF    THE    COMMON    METALS.  405 

Deducting  treatment,  $4+  (4SiO2  excess  by  loc) 

+  5oc  zinc  penalty 4 . 90 


Net  returns  to  the  shipper  f.  o.  b.  Denver $39-59 

On  the  flat  schedule 

Gold  0.2  oz.  @  $19.50 $  3.90 

Silver  60  oz.  @  95%  of  62c 35-34 

Lead  21%  @  2$c 5.25 


Gross  metal  values    $44.49 

Deducting  treatment  $6.50+500  zinc  penalty 7.00 


Net  returns  to  the  shipper  f.  o.  b.  Denver $3749 

Therefore,  the  neutral  basis  figures  better  for  the  shipper,  and 
is  the  one  used.  Had  both  silica  and  lead  been  somewhat  more, 
the  flat  schedule  would  have  proved  more  profitable  for  the 
shipper. 

125.     THE  MARKETING  OF  ORES  AND  METALS. 

The  metallurgist  must  pay  particular  attention  to  the  kind  of 
ton  used  in  weighing  ores  and  the  common  metals,  and  not  take 
the  value  for  granted  without  a  clear  understanding  of  condi- 
tions. The  short  ton  of  2,000  Ib.  is  used  in  the  Western  States 
for  ores  and  metals.  In  the  Eastern  States  the  long  ton  of  2,240 
Ib.  is  used  for  coal,  iron  ore  and  pig  iron.  In  England,  copper, 
spelter  and  tin  are  weighed  by  this  ton.  We  may  also  note  that  the 
metric  ton  of  1,000  kilos  or  2,204  Ib.  approaches  the  long  ton  in 
weight. 

New  York  is  a  chief  market  for  metals,  and  the  sales  of  ores 
and  metals  are  based  upon  these  prices.  The  quotations  of  other 
markets  as  San  Francisco  and  London  are  also  often  given. 
Referring  to  a  technical  periodical,  as  for  example  the  Mining 
and  Scientific  Press  of  San  Francisco  or  the  Engineering  and 
Mining  Journal  of  New  York,  we  find  the  following : 

Silver.  At  New  York  the  quotations  are  per  troy  ounce  of 
fine  silver,  1000  fine.  It  takes  14.58  ounces  troy  to  make  a  pound 
avoirdupois.  London  prices  are  for  sterling  silver  925  fine. 


BANCROFT 


406  THE    METALLURGY 

The  exchange  value  of  English  or  sterling  money  is  also  given. 
Thus  it  is  possible  to  reckon  the  foreign  price  of  metals-  in  dol- 
lars and  cents. 

Copper.  At  New  York  the  value  of  this  metal  is  expressed 
in  cents  per  pound,  quotations  being  given  for  Lake  copper  in 
cakes  for  rolling  into  sheets,  ingots  for  re-melting  to  make  cast- 
ings and  brass,  and  wire  bars  for  drawing  into  wire.  Cathodes 
or  cathode  plates  are  of  copper  in  rough  sheets,  the  product  of 
electrolytic  refining.  When  these  cathodes  have  been  re-melted 
and  cast  into  cakes,  ingots  or  wire  bars,  the  metal  is  called  elec- 
trolytic. It  sells  at  about  %  cent  per  pound  less  than  Lake  cop- 
per. Cathodes  cost  *4  cent  per  pound  less  than  electrolytic  copper 
because  it  nominally  costs  that  much  to  re-melt  them.  Casting- 
copper  is  a  lower  grade,  not  so  pure,  but  well  suited  for  making 
brass  castings. 

At  London,  copper  is  sold  by  the  long  ton  in  English  money, 
and  is  of  various  grades,  which,  with  a  sample  set  of  prices,  is 
as  follows : 

English  tough  copper £70.  los  @   £71.  os 

Best  selected    £71.  155  @   £72.  55 

G.  m.  b's  (Good  merchantable  bars)  or  stand- 
ard     £66  155  spot 

£66  175  6d.,  3  mo? 

This  latter  quotation  has  reference  to  whether  the  copper  is 
for  immediate  delivery  or  whether  the  customer  will  be  ready  to 
take  it  at  the  expiration  of  three  months,  by  which  time  the  re- 
duction works  may  have  produced  it. 

Strong  sheets  (rolled  copper) £79  los 

India  sheets   (for  sheeting  vessels) £75   los 

Yellow  metal  (a  grade  of  brass) 6^d  per  Ib. 

It  is  the  business  of  dealers,  and  others  interested  in  copper, 
to  keep  statistics  of  the  copper  on  hand,  which  is  called  the 
visible  supply.  When  there  is  but  little  copper  on  hand  the 
price  naturally  goes  up  and  vice  versa. 

Tin. — Like  copper,  tin  may  be  quoted  for  immediate  or  for 
future  delivery  at  a  specified  time.  A  sample  of  a  quotation 
would  be  3O^Jc  for  spot,  29^  @  3oc  for  futures.  The  prices  are 


OF    THE    COMMON    METALS.  407 

die  same  whether  the  metal  is  from  the  Burmah  near  the  Straits 
of  Malacca  (Straits  tin)  from  Bolivia,  from  Australia,  or  from 
other  countries. 

Zinc  is  called  commercially  spelter.  Quotations  in  the  United 
States  are  given  in  cents  per  pound  thus : 

New  York  5.80®  5.850. 

St.  Louis  5-65@5-7o~c. 

St.  Louis  is  close  to  the  zinc-producing  district  of  Kansas  and 
Missouri,  and  hence,  the  lower  price  for  spelter  as  compared 
with  New  York.  In  the  London  market,  spelter  is  quoted  by  the 
long  ton.  Thus  we  have  a  sample  quotation  £24  for  good  ordi- 
naries (ordinary  grades),  and  £24  55  for  special  brands  or 
makes.  Prices  for  the  ore  are  often  given  in  tons  of  2000  Ib. 
at  Joplin,  Mo.  They  make  a  basis  price  for  zinc  ore  assaying 
60%  zinc  and  vary  this  according  to  the  zinc  contents.  It  used 
to  be  the  custom  to  figure  the  price  of  basis  ore  per  ton  in  dollars 
by  multiplying  the  price  of  the  zinc  per  pound  by  7.5  but  this 
proportion  is  only  an  approximation,  buyers  often  advancing  their 
prices  to  secure  ore  when  it  is  scarce. 

Mercury  or  quicksilver. — A  sample  quotation  may  be  given 
thus:  Quicksilver  is  easy  at  $38@$35-5o  per  flask  in  large  lots, 
and  $40  on  smaller  orders.  The  London  price  is  £7.  I2s.  6d. 
at  second  hand,  that  is,  sold  by  jobbers  and  not  by  the  reduction 
works.  A  flask  used  to  weigh  86%  Ib.  but  this  has  been  changed 
so  that  a  flask  must  contain  75  Ib.  of  mercury. 

Precious  metals. — Gold  is  sold  to  the  mints  at  a  steady  price 
of  $20.67  Per  troy  ounce.  This  price  is,  however,  subject  to  a 
deduction  of  about  2c  per  ounce  for  melting  and  assaying.  Plat- 
inum though  a  commercial  metal  is  of  about  the  same  value  as 
gold  say  $20.50  per  ounce.  Silver  has  already  been  mentioned. 


T.  A.  RICKARD,  Editor 


ESTABLISHED  MAT  24,  1860.  ll»\    >* 

EDGAR  RICKARD,  Bus.  Mgr. 


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